IC 8846 



Bureau of Mines Information Circular/1981 

c 2 






Pumped-Slurry Backfilling 

of Abandoned Coal Mine Workings 

for Subsidence Control 

at Rock Springs, Wyo. 

By G. J. Colaizzi, R. H. Whaite, and D. L. Donner 






(^r^>s 



UNITED STATES DEPARTMENT OF THE INTERIOR 



Information Circular 8846 



Pumped-Slurry Backfilling 

of Abandoned Coal Mine Workings 

for Subsidence Control 

at Rock Springs, Wyo. 

By G. J. Colaizzi, R. H. Whaite, and D. L. Donner 

With an appendix on Hydraulic Model Studies for Backfilling Mine 
Cavities by E. J. Carlson, U.S. Bureau of Reclamation, Denver, Colo. 




UNITED STATES DEPARTMENT OF THE INTERIOR 
James G. Watt, Secretary 
BUREAU OF MINES 




As the Nation's principal conservation agency, the Department of the Interior 
has responsibility for most of our nationally owned public lands and natural 
resources. This includes fostering the wisest use of our land and water re- 
sources, protecting our fish and wildlife, preserving the environmental and 
cultural values of our national parks and historical places, and providing for 
the enjoyment of life through outdoor recreation. The Department assesses 
our energy and mineral resources and works to assure that their development is 
in the best interests of all our people. The Department also has a major re- 
sponsibility for American Indian reservation communities and for people who 
live in Island Territories under U.S. administration. 




3c~&%ini 



This publication has been cataloged as follows: 



Colaizzi, Gary J 

Pumped-slurry backfilling of abandoned coal mine workings 
for subsidence control at Rock Springs, Wyo. 

(Bureau of Mines information circular ; 8846) 

Bibliography: p. 55*56. 

Supt. of Docs, no.: I 28.23:8846. 

1. Mine subsidences— Wyoming— Rock Springs. 2. Mine filling— Wyo- 
ming— Rock Springs. 3> Coal mines and mining— Environmental aspects- 
Wyoming— Rock Springs. 4. Slurry. I. Whaite, Ralph H., joint author. 
II. Donner, Donald L., joint author. III. Title. IV. Series: United 
States. Bureau of Mines. Information circular ; 8846. 



TN295.U4 [TN319] 622s [622'.334] 80-39876 



For sale by the Superintendent of Documents, U.S. Government Printing Office 

Washington, D.C. 20402 



CONTENTS 



Page 



Abstract 1 

Introduction 1 

Purpose and scope of report 4 

Acknowledgments 4 

Background 5 

Subsidence problems 5 

Subsurface conditions 7 

Hydraulic backfilling methods 10 

Controlled flushing 10 

Blind flushing 10 

Pumped-slurry injection 11 

Pumped-slurry test at Rock Springs 15 

Inj ection operation 16 

Evaluation 18 

Cost 19 

Reports 20 

First large-scale test 20 

Hydraulic model studies for backfilling mine cavities 21 

Three large-scale projects at Rock Springs 21 

Drilling operations 22 

Slurry components ,. 24 

Mixing plants and slurry pumps . . 26 

Project no. 1 27 

Project no. 2. . .. 33 

Project no. 3, with a management support contract 43 

Conclusions 51 

References 55 

Appendix. --Hydraulic model studies for backfilling mine cavities, by 
E. J. Carlson (follows references) 

ILLUSTRATIONS 

1. General location map of Rock Springs, Wyo 6 

2. Map of Rock Springs showing the 15 critical areas 8 

3. Typical section of the coal seams under the city 9 

4. Typical sections showing boreholes and configuration of fill mate- 

rial placed in mine voids by the gravity blind flushing method. ... 10 

5. Typical residential block showing the pattern of boreholes in 

streets, alleys, and vacant lots proposed for blind flushing 12 

6. Sectional views through a flooded mine room at the point of slurry 

injection showing movement of particles and growth of deposit 13 

7. Top view of mine model with transparent roof, showing stages of 

radial distribution of fill material by the pumped-slurry process. 14 

8. Photograph of the plant site at the first test of the pumped-slurry 

mine backfilling technique 16 

9. Plan view of installed equipment at the plant site and adjoining 

streets 17 



ii 



ILLUSTRATIONS --Continued 

Page 

10. Map of mine workings at the site of the first demonstration project 

in Rock Springs, Wyo. , showing pattern of sand deposition in the 

mine voids 19 

11. Typical sections of water supply and injection boreholes 23 

12. Typical section of a monitoring borehole 23 

13. Map of a portion of Rock Springs showing locations of critical 

areas 1, 7, and 8 in relation to mine workings of the No. 7 Seam. 28 

14. View of the slurry mixing plant installation at the site of the 

borrow pit about 2 miles from the built-up areas of the city 30 

15. Map showing location of water-supply wells and pipeline, slurry 

pipelines, and injection boreholes in relation to critical 

areas 1 , 7 , and 8 31 

16. Map of a portion of Rock Springs showing location of critical 

area 14 and a part of critical area 11 in relation to underground 
mine workings in the No. 1 Seam 34 

17. Map of the project area showing location of the slurry plant in 

relation to water-supply wells and pipeline, slurry pipelines, 

and injection boreholes 36 

18. View of earth -moving and screening equipment operations in the 

borrow pit 38 

19. View of end of hopper, conveyor, slurry pump, diesel engine, pipe- 

lines and control building at mixing plant 38 

20. View of mixing tank, suction line, sand being delivered to mixing 

tank, and bulldozer in the background pushing sand into hopper. . . 39 

21. View of the installed slurry pipeline in the pedestrian underpass 

bene'ath the mainline tracks of the Union Pacific Railroad 40 

22. Map of a large portion of Rock Springs showing location of critical 

area 2 in the No. 7 Seam and Critical areas 9, 10, 11, 12, 13, 

and 15 in the No. 1 Seam in relation to underground mine workings 44 

23. Map of a portion of Rock Springs showing the location of Project 3 

ins tal la t ions 46 

24. View of earth-moving and screening equipment being used in the 

borrow pit 48 

25. Photograph of sand being bulldozed into hopper and fed from con- 

veyor into mixing tank 48 

26. View of mixing plant operation 49 

27. View of slurry pipeline on the pedestrian foot bridge over Bitter 

Creek 50 

TABLES 

1. Data on solids, water, and slurry 26 

2. Injection data on the initial large-scale demonstration project, 

1973-74 32 

3. Injection data on the second large-scale demonstration project, 

1974-75 42 

4. Injection data on the third large-scale demonstration project, 1976 50 



PUMPED-SLURRY BACKFILLING OF ABANDONED COAL MINE WORKINGS 
FOR SUBSIDENCE CONTROL AT ROCK SPRINGS, WYO. 

by 
G. J. Colaizzi, 1 R. H. Whaite,2 and D. L. Donned 



ABSTRACT 

The Bureau of Mines, at the request of local authorities in Rock Springs, 
Wyo. , investigated and conducted through contracts a multistage program of 
exploratory drilling and pumped-slurry backfilling of 15 areas of potential 
subsidence in abandoned mine workings underneath that community. Initially, 
the Bureau in 1969 had recommended a program of gravity blind flushing of some 
of the inaccessible mine voids, and in 1970 a new technique, the pumped-slurry 
injection process, was tested for the first time in a site adjacent to the 
city's area of severe surface subsidence. Success of this initial .testing 
program, and of a large-scale project in Scran ton, Pa. , led to further large- 
scale projects, funded by Congress, that resulted in the successful back- 
filling not only of all 15 target areas of potential subsidence in Rock 
Springs, but also of several areas in other States. Total cost of the proj- 
ects in Rock Springs, including the original pumped-slurry test, was 
$3,243,993. A total of about 923,000 tons of sand was injected hydraulically 
into mine voids, rendering 178 acres of residential and central -downtown areas 
of Rock Springs less susceptible to subsidence damage. The pumped-slurry 
method was proved to be much superior to the gravity blind flushing method in 
terms of the amount of solids that could be injected underground through a 
single borehole. However, there are special conditions that make this tech- 
nique more or less applicable in different areas or underground configurations, 
as noted in the report's conclusions. 

INTRODUCTION 

Surface subsidence is often the consequence of underground mining opera- 
tions. Removal of solid material from beneath the earth's surface produces 
voids, and once the natural support afforded by this material is taken away, 
the weight of the overburden is redistributed. If pillars of material that 
are left unmined or timbers or other artificial support left underground are 

Supervisory mining engineer. 
3 Mining engineer. 
3 Research supervisor. 

All authors are with the Mine Environment and Reclamation Division, Denver 
Research Center, Bureau of Mines, Denver, Colo. 



not sufficiently strong to support the overburden, the overlying rock breaks 
and falls into the voids and/or crushes the pillars. This breakage may pro- 
ceed upward in overlying material as far as the surface, causing potholes, 
cracks, or general settling of the ground. The time between the completion 
of mining operations and the disturbance at the surface may be a matter of 
days or a period of many years, depending upon a number of factors including 
the nature of the overlying rocks, depth of the excavation beneath the surface, 
and the method of mining employed. Subsidence is postponed when correctly 
spaced and adequate-size pillars are used to provide overburden support or is 
minimized when voids are backfilled with suitable material for the same 
purpose. 

In the United States, nearly 100,000 underground mines are in existence, 
of which an estimated 90,000 are closed or abandoned. The total land that has 
been undermined for the production of coal, metals, and nonmetallic minerals 
has been estimated to be about 7-1/2 million acres. The exact percentage of 
the undermined land that has been affected by subsidence is not known. In a 
recent study of land utilized by the mining industry, the surface area that 
had subsided or was otherwise disturbed by underground mining from 1930 
through 1971 was estimated at 105,000 acres. Of this acreage, 84 percent 
resulted from mining coal (15) . 4 

Most of the land affected by subsidence is removed from centers of popu- 
lation. Adverse effects in these areas consist of crop damage, altered drain- 
age patterns, and reduced land values. The most severe damage, of course, has 
occurred in urban areas. Millions of dollars in damage has resulted from the 
differential settling of buildings, pavements, subsurface pipelines, and other 
facilities, compounding existing problems as the pressure to develop under- 
mined land is increasing in many metropolitan areas in response to accelerat- 
ing demands for more living space. 

For the control of subsidence in built-up areas where the underlying 
mines have been abandoned, backfilling is believed to provide the most prac- 
tical means of minimizing damage to the communities. From time to time, 
studies of subsidence problems have been made for the city of Scranton (_7, 11 , 
23 ) and for the Commonwealth of Pennsylvania (16-17) . All the studies recom- 
mended programs of hydraulic backfilling. 

The first reported use of hydraulic backfilling of mine workings was in 
the Anthracite region over 100 years ago. The purpose was to stop the sub- 
sidence of a church, and the treatment succeeded (2, pp. 99-100). Hydraulic 
backfilling was developed during the late 1800' s and early 1900' s and was used 
in about one-fourth of the anthracite mines for such purposes as to extinguish 
mine fires, to arrest the development of progressive pillar failure known as 
mine squeeze, to permit the reclaiming of pillars, to dispose of unwanted mine 
refuse, and to protect the surface. The practice of backfilling by the coal 
industry in the United States decreased after World War I with the decline of 
the anthracite industry. In domestic bituminous coal mining operations, 

4 Underlined numbers in parentheses refer to items in the list of references 
preceding the appendix. 



backfilling has never been common practice (4). Applications to metal mining, 
however, in the United States and elsewhere, have provided solutions to a 
variety of ground control problems resulting in greater resource recovery, 
safer working conditions, and reduced mining costs (12) . 

The principal development of hydraulic backfilling in coal mining took 
place in Europe in the early 1900' s, where the practice had spread from the 
Anthracite region of the United States. Most European coalfields differ from 
those in the United States in that the coalbeds are much deeper, the concen- 
tration of coal within a vertical section is much greater, and longwall mining 
methods predominate, whereas room-and-pillar mining is the most common method 
in the United States (18, p. 35). Many European coal mines underlie highly 
developed industrial areas or commercial waterways that require protection. 
The purposes of backfilling are to contribute to roof control under ground 
pressures and to permit as nearly complete recovery of the coal as possible 
as well as to support the surface. Hydraulic backfilling remains a part of 
mining operations in some European coalfields where thick beds of coal are 
being removed from beneath densely populated areas such as in France and 
Poland. 

The Bureau of Mines 1 interest in hydraulic backfilling is as old as the 
Bureau itself. The First Annual Report of the Director (8, pp. 42-43) des- 
cribes an ongoing study of mine filling (hydraulic backfilling) to make the 
mines a safer place in which to work and to reduce settlement of the surface. 
The report resulting from that study covered the history, applications, meth- 
ods, and costs of hydraulic backfilling (6) . Other early Bureau publications 
reported on sources of backfill materials and applications of hydraulic back- 
filling to various mining problems. At the end of World War II, the Bureau 
looked into the backfilling problem as it related both to the conservation of 
anthracite and to the prevention of subsidence in order to determine what role 
the Federal Government might play (I) . A comprehensive engineering study of 
the backfilling problem in the Anthracite region by the Federal Government was 
recommended; the work was to be done in cooperation with the Commonwealth of 
Pennsylvania, and with the anthracite industry, should the study lead to 
action. 

The current participation by the Bureau of Mines in subsidence-control 
projects in areas of abandoned mines is provided for by two pieces of legis- 
lation that authorize Federal-State cooperation. Public Law 87-818, an 
amendment in 1962 to the 1955 Anthracite Mine Drainage Act (Public Law 84-162) , 
authorized the Secretary of the Interior to participate equally with the Com- 
monwealth of Pennsylvania in the filling of voids in abandoned anthracite 
mines, in those instances where such work is in the interest of the public 
health or safety. Under this provision, four subsidence-control projects were 
completed between 1962 and 1965. The Appalachian Regional Development Act of 
1965 (Public Law 89-4) and its amendments of 1967 (Public Law 90-103) included 
authorization to fill voids in abandoned coal mines within the Appalachian 
region. Costs under the Appalachian Act are shared 75 percent by the Federal 
Government and 25 percent by the cooperating State. By the end of 1978, fif- 
teen subsidence control projects had been completed under the Appalachian pro- 
gram and additional projects were in progress in the Anthracite region of 



Pennsylvania. Under the same program one project in Maryland and one in West 
Virginia have also been completed. 

Under the authority assigned to the Bureau of Mines by the Organic Act 
(May 16, 1910) and its succeeding amendments and pursuant to regulations 
(30 U.S. Code 1-11), the Bureau conducts scientific and technologic investi- 
gations concerning mining and its related problems. By the end of 1978, 
twenty-one subsidence-control demonstration projects had been conducted under 
this authority, including the Rock Springs, Wyo. , projects and the demonstra- 
tions in Pennsylvania, Maryland, West Virginia, and Illinois. 

In demonstration projects conducted under the Organic Act, the Bureau is 
the primary signatory to any contract that is entered into. Aside from any 
agreements with contractors, the Bureau enters into cooperative agreements 
with State or local authorities. Through these agreements the State or local 
government, referred to as the "cooperator," grants the Bureau and its con- 
tractors the right to enter onto public streets and land in order to carry out 
the proposed project work. 

A Draft Environmental Impact Statement, DES 75-37, on "Surface Subsidence 
Control in Mining Regions," was used (prior to its final approval in November 
1976) as a guide in conducting the subsidence control projects in Rock Springs, 
Wyo. (21). 

Purpose and Scope of Report 

This report reviews the investigation by the Bureau of Mines to determine 
the cause and magnitude of the subsidence problem at Rock Springs, Wyo. It 
discusses methods of backfilling mine voids that are used to minimize the 
effects of subsidence. Four subsidence control demonstration projects at Rock 
Springs are discussed in detail, and an assessment is made of the effective- 
ness of the work performed. 

A cknowl ed gmen t s 

The contributions of many individuals to this report, and to the demon- 
stration projects on which it is based, are gratefully acknowledged. Paul L. 
Russell, former Research Director of the Bureau's Denver Research Center, was 
responsible for Bureau of Mines activities at the demonstration projects in 
Rock Springs. Neil Morck, District Manager for the U.S. Bureau of Land 
Management at Rock Springs, provided leases for two of the borrow pits and 
stipulations governing their use. Robert Oster, District Engineer for the 
Union Pacific Railroad, provided leases for use of the railroad's right of 
way. 

Officials of the city of Rock Springs, including Paul Wataha, Mayor, 
Wayne Johnson, City Engineer, and Hyram Fedji, Director of Planning, Engineer- 
ing and Development, provided excellent cooperation in the performance of the 
work on city streets. Valuable information to the drillers on the location of 
utilities under streets was obtained from Mountain States Telephone and Tele- 
graph companies and the Mountain Fuel Supply Co. Data on surface altitudes 



from several bench marks in the three large-scale project areas were made 
available to the Bureau as needed by the firm of Johnson-Fermelia & Crank, Inc. 

Injection operations at the first test of the pumped-slurry process and 
at the first large-scale demonstration, both at Rock Springs, were performed 
by the Dow Chemical Co. , Dowell Division, of Tulsa, Okla. A concept that sand 
slurry could be transported in pipelines directly from the borrow pit to the 
injection boreholes, a distance of over 2 miles, demonstrated additional 
flexibility in the use of the pumped-slurry process. Dowell Division person- 
nel who contributed to the success of these nonroutine operations were 
John D. Stewart, Milton E. Heslep, Robert Hurst, L. D. Boughton, and George 
Laflin; and also A. J. Meyers on the use of sonar caliper surveys. 

Engineers of the WHAN Co. and of the Bober Co. , contractors for the 
second and third large-scale projects, contributed to improving the overall 
operations through more efficient use of the automated controls of the slurry 
mixture. In this way, delivery of a more uniform mixture to the injection 
boreholes was gradually accomplished with a considerable saving in the use of 
power. As the work progressed they relocated the equipment within the con- 
fines of the plant site and altered the use of the equipment from time to time 
to improve performance and eliminate as much noise and dust as possible. 

Charles S. Kuebler, Chief of the Bureau's Environmental Affairs Field 
Office at Wilkes-Barre, Pa. , where similar type projects were being conducted, 
and E. J. Carlson, engineer, Hydraulics Branch, U.S. Bureau of Reclamation, 
Denver, Colo., in charge of the model studies, provided valuable insight into 
the operation of the pumped-slurry method and the behavior of the slurry. 

Much of the description of the nature of mine subsidence, its history, 
and various methods of control, including discussions of the pumped-slurry 
technique that appear in this report, is attributed to Alice S. Allen and 
Ralph H. Whaite, authors of Bureau of Mines Information Circular 8667 (22) . 

BACKGROUND 

Development of Rock Springs, Wyo. , as a mining community began in the 
1860 's with the westward extension of the Union Pacific Railroad through south- 
ern Wyoming. The railroad's demand for coal persisted well into the 20th cen- 
tury, leaving about 900 acres of Rock Springs undermined. Extraction of coal 
by the room and pillar method under most of the city had reached an advanced 
stage, and according to available mine maps it appeared that many of the 
remaining pillars were too small to support the overburden indefinitely. 
Under other areas some of the pillars had been removed. Reports of surface 
subsidence in various parts of the city became noteworthy in the late 1960's. 
Damage to houses, commercial buildings, streets, gas mains, waterlines, and 
sewers was considerable. 

Subsidence Problems 

The city of Rock Springs (fig. 1) is located in Sweetwater County in 
southwestern Wyoming on U.S. Highway 30, Interstate 80, and the Union Pacific 



r 



MONTANA 



.Sheridan 



Buffalo* 



W Y M I 



(789) 



Rawlins^ 



7 3 ON 



Rock Springs 



1807 



Green River 




Casper 



COLORADO 



30 



Scale, miles 



60 

_J 



FIGURE 1. - General location map of Rock Springs, Wyo. 

Railroad. The 1969 population of the city was about 12,000. Mining of bitu- 
minous coal had been the principal industry in the city from the 1860's to the 
1950' s, and most of the built-up area is extensively undermined, a condition 
typical of many of the early mining communities. As a consequence, subsidence 
of the surface, locally termed "sink holes," has plagued Rock Springs for many 
years. Both the State Highway Department and the Union Pacific Railroad have 
filled, in some areas, abandoned mine workings beneath their respective rights 
of way. 



In 1969, the Bureau of Mines conducted an investigation of subsidence in 
Rock Springs, at the request of local authorities, to determine the cause and 
to recommend solutions (5) . Severe subsidence damage had been experienced 
since 1967 within a 2-acre area in the eastern portion of the city, affecting 
at least 18 houses and damaging streets, sidewalks, gas mains, waterlines, 
and sewers. Subsidence was gradual and continuing, achieving maximum 



settlement of about 30 inches, accompanied by lateral displacements and some 
heaving. Elsewhere in the city, localized "potholes" continued to appear on 
the surface of the ground from time to time. 

The site of the severe damage was about 6 blocks from the center of the 
city near the intersection of Connecticut Avenue and D Street. About one-half 
of the damage at the time of the investigation was to property known as the 
Rock Springs Camp owned by the Mountain Fuel Supply Co. The camp consisted of 
seven dwellings and one office building, all on the north side of D Street. 
Individual private homes on the south side of D Street were also affected. 
Ten of the 18 affected houses were seriously damaged. 

Subsurface Conditions 

From a study of available mine maps it was determined that collapse over 
mine openings was the principal cause of the subsidence in the Connecticut 
Avenue-D Street area and in other parts of the city as well. Of the 900 acres 
of the city that was undermined, Bureau engineers reported that in 15 separate 
areas ranging in size up to 46 acres, the conditions in the mines with respect 
to supporting pillars were comparable to those under the Connecticut Avenue- 
D Street area. These critical areas comprising about 210 acres constituted a 
significant part of the city, and facilities therein served a wide range of 
functions. They include sections of the central business district, the city 
hall, and 828 dwelling units. At the time of the investigation in 1969, it 
was estimated that the total value of the houses alone was nearly $11 million. 
Figure 2 is a map of Rock Springs showing the locations of the 15 critical 
areas. 

Geologically, the city of Rock Springs is located on the west edge of the 
Rock Springs Uplift (10, 20) . Bedrock in the area is the Rock Springs Forma- 
tion of Cretaceous age--an irregular series of coalbeds, carbonaceous shale, 
siltstone, claystone, and sandstone. The strike of the formations is N 36° E, 
and the dip is 6° to the northwest at the site of the subsidence. Downdip 
(northwest of the site) the dip steepens to 30°. Faults are common in the 
area, but under the built-up section of the city, faulting is believed to be 
of relatively minor consequence. Figure 3 is a typical cross section of the 
coal seams under the city of Rock Springs. 

Depth to bedrock ranged from 6 to 52 feet in boreholes that were drilled 
in the Connecticut Avenue-D Street area. The overlying alluvium is silty, 
very fine sand. A deposit of clay beneath the silty sand was reported in one 
borehole. The silty sand deposit is of the type that is subject to compaction 
in the presence of excess water, hence a potential contributor to subsidence. 
However, in the Rock Springs subsidence area, the evidence indicated that col- 
lapse over mine openings was the principal cause of subsidence (13). Deeper 
strata within the bedrock sequence were found to have been displaced downward, 
and the pattern of subsidence at street level was observed to reflect the 
spacing of rooms and pillars in the mine below. 




FIGURE 2. - Map of Rock Springs showing the 15 critical areas. 



V) 

■t- 

E 



Looking N 36 E 



'e 




O 1000 
Scale, feet 

FIGURE 3. - Typical section of the coal seams under the city (note distorted scale; actual dip 
of seams is 6°). 

In the vicinity of Rock Springs, the aggregate thickness of coal exceeds 
90 feet. Under the built-up part of the city, two coalbeds were mined exten- 
sively, but their structural position is such that one bed has been mined 
beneath another mined bed in only one section of the city. Two other coalbeds 
were mined to a lesser degree. The room-and-pillar system of mining was used 
exclusively. Patterns of extraction, however, were irregular owing to minor 
faulting and changes in mining techniques over the years. 

Water presented a problem during mining operations. Reportedly, one mine 
pumped 500,000 gallons per day when operating, and water-related problems were 
a factor in the closing of another mine in the demonstration area. Water lev- 
els measured in drill holes indicated that about 75 percent of the workings 
are now flooded. 

A key element in planning successful backfill operations is the availabil- 
ity of accurate and complete mine maps. The information available from mine 
maps on coalbeds beneath Rock Springs was not complete, and data obtained from 
drilling indicated that the maps would require adjustment to correlate with 
surface maps. About a third of the exploratory boreholes, which were drilled 
to intersect openings according to the maps, terminated in solid pillars of coal. 



10 



The Bureau of Mines report included a recommendation that the mine voids 
in the 15 critical areas, most of which were flooded, should be filled with 
sand to support remaining pillars, reduce rock strata breakage in the over- 
burden, and minimize possible damage at the surface. To backfill the entire 
undermined area of the city, about 900 acres, would be difficult and quite 
expensive because most of the voids, exclusive of the critical areas, are 
situated at depths in excess of 300 feet* below the surface. Based on local 
observation and experience gained in subsidence control in the Anthracite 
region, it was assumed that strata breakage originating at depths below 
300 feet would not be significant at the surface. 

HYDRAULIC BACKFILLING METHODS 

Before 1970, two methods of hydraulic placement of backfill material 
in underground mined-out spaces had been used in the Anthracite region. They 
are known as controlled flushing and blind flushing. In both methods, granu- 
lar solid material is sluiced down from the surface through boreholes with 
water by the force of gravity. 

Controlled Flushing 

Controlled flushing is possible in mines in which men can safely enter 
and gain access to key areas for the filling operations. Bulkheads are built 
in mine passages around the periphery for containment of the fill. Drain 
boxes may be incorporated in these structures to facilitate rapid removal of 
water. The injection boreholes, generally about one borehole for 4 acres, are 
cased from the surface to the mine opening. At the base of each hole, long- 
radius 90° pipe elbows are placed through which slurry is diverted to horizon- 
tal pipes and distributed into the mine workings; Horizontal dispersal ranges 
from 300 to 1,000 feet, depending on the vertical distance from the ground 

surface to the mine opening 
Boreholes 



Surface 




Surface- 




FIGURE 4. - Typical sections showing boreholes and con- 
figuration of fill material placed in mine 
voids by the gravity blind flushing method. 



and the solids concentration 
of the slurry. Controlled 
flushing provides the best 
support and is used where con- 
ditions permit. 

Blind Flushing 

Many abandoned mine open- 
ings are inaccessible because 
of flooding or extensive cav- 
ing. Such openings were 
flushed blindly in the past. 
The gravity- feed method simply 
builds a conical pile beneath 
the underground opening of the 
flushing hole. When the apex 
of the cone builds up to the 
mine roof, no more fill will 
enter the mine opening. Fur- 
ther injection must use other 
boreholes, as shown in figure 4. 



11 



Depending on conditions underground, such as the dip of the bed, its height, 
and the proximity of pillars or occasional caved roof strata, the volume of 
material that can be injected in the this manner from each borehole ranges 
from 20 to 1,000 cubic yards. In a 6-foot bed that is relatively flat, for 
example, in which about 45 percent of the bed remains in pillars, only about 
100 cubic yards can be injected from a single hole. Therefore, injection 
holes must be closely spaced, but at best only about a third of the under- 
ground open space is filled by this blind flushing technique. Most blind 
flushing projects have required hundreds of flushing holes. In built-up areas, 
it may not be possible to drill boreholes in critical areas where buildings or 
other structures interfere or where easements cannot be obtained from property 
owners. Therefore, most of the backfilling under built-up areas is done 
through boreholes drilled in streets and alleys, and the support given is of 
only indirect benefit to adjoining buildings (fig. 5). 

Pumped-S lurry Injection 

A new technique for the blind flushing of inaccessible mine workings, 
termed "the pumped-slurry injection process," was first demonstrated in 1970 
at Rock Springs, Wyo. , in a test site adjacent to the Connecticut Avenue- 
D Street area where severe subsidence problems had developed. This technique 
differs from the open gravity-feed methods previously described in that pump- 
ing energy is used to achieve a dynamic suspension of solid particles in water 
and the system is completely closed from the point of slurry mixing to the 
bottom of the injection borehole. 

In this process, granular material is blended with water, and the sus- 
pension (slurry) is pumped to the point of deposition. Water from an inun- 
dated mine may be used and recirculated, or water from an external source may 
be used without being recirculated. During mixing, each solid particle 
becomes enclosed by fluid so that friction during transit is reduced. The 
slurry is pumped continuously from the mixing tank through a pipeline on the 
surface and thence down through a borehole to the mine opening. The energy 
provided by the pump and the static head in the borehole give the velocity 
required to keep the solid particles in suspension and to transport them. 

The completeness of filling in the open spaces is responsive to changes 
in the velocity of flow, which changes with the growth of the mound of depos- 
ited solids. As the slurry first enters the open space from the injection 
hole, its velocity drops rapidly, and solid particles settle out near the bore- 
hole, forming a doughnut-shaped mound on the mine floor. As the height of the 
mound approaches the mine roof, the velocity of the slurry increases through 
the narrowing channel, and solid particles are transported to the outer limit 
of the mound. Here the velocity decreases abruptly and solids are deposited. 
This type of deposition continues and the mound of deposited fill builds 
outward. Stages in the filling of a mine void are shown diagrammatical ly 
in figure 6. 



12 



^ \>> 




O Borehole 

^| Fill material 

S* House 

FIGURE 5. - Typical residential block showing the pattern of boreholes in streets, 
alleys, and vacant lots proposed for blind flushing. Circular areas 
around boreholes represent backfill material to be placed in the mine 
voids. 



13 



PUMPED SLURRY 



r 



INJECTION BOREHOLE 



"►ft 




FLOODED MINE ROOM 




ttUWUUM 



ROCK STRATUM 




DEPOSITED SOLIDS 




FIGURE 6. - Sectional views through a flooded mine room at the point of slurry injection 
showing movement of particles and growth of deposit. 



14 









FIGURE 7. - Top view of mine model with transparent roof, showing stages of radial dis- 
tribution of fill material by the pumped-slurry process. 

{Courtesy of the Dowell Division of the Dow Chemical Co.) 



15 



In a table-size mine model simulating the arrangement of rooms and pil- 
lars, deposition of fill material is shown in figure 7. As resistance to flow 
of the slurry develops in one direction, a new channel is formed in another 
direction along a line of less resistance. Eventually, nearly all mine open- 
ings are filled. The lateral extent of the fill is determined largely by the 
available energy in the system. As the mound of fill material builds outward 
in the mine, the flow channels between the mound and the mine roof become 
longer and resistance to flow increases. When this resistance, combined with 
resistance in the pipe, becomes great enough to reduce the velocity of the 
slurry below that required to transport the solid particles, transportation 
of the particles ceases. The particles then settle out, and the passage 
becomes filled. 

The pumped-s lurry method of blind flushing has the following advantages 
over the open gravity-feed method previously used: 

1. Great reduction in the number of injection boreholes. A single injec- 
tion hole serves the purpose of many injection holes in the gravity-feed 
method. 

2. More complete vertical filling of mine openings. 

3. More complete areal coverage. Areas inaccessible because of surface 
improvements can be filled. 

4. Less disruption of the community in the form of noise, dust, and 
traffic interference by drilling operations and trucking of fill material. 

Disadvantages to the use of the pumped- slurry method would be apparent 
under the following conditions: 

1. Where it is safe for men to enter the mine and control the backfill- 
ing operation. 

2. Where the mine voids to be filled are close to the surface with lit- 
tle or no rock strata cover, in which case the slurry water could invade 
surface areas. 

3. Where the amount of mine void to be filled is too small to warrant 
the cost of acquiring and installing slurry pumping equipment. 

PUMPED-SLURRY TEST AT ROCK SPRINGS 

The test at Rock Springs in 1970 was carried out in cooperation with the 
city of Rock Springs, the Department of Housing and Urban Development, and the 
Bureau of Mines. The Bureau's financial participation in the project to the 
extent of $55,000 was covered by a contribution contract with the city. The 
city of Rock Springs entered into a contract with Dowel 1 Division of the Dow 
Chemical Co. of Tulsa, Okla. , to perform the injection operation. The total 
contract cost of the demonstration was $173,140. 



16 



A site adjacent to the Connecticut Avenue-D Street area was selected for 
the test because it was believed to have a high subsidence potential. Explor- 
atory drilling in conjunction with sonar scanning indicated the mine workings 
under the site were flooded and included both collapsed areas and open spaces 
in the coalbed. Open passages usable both for water supply and receiving a 
large quantity of fill material were identified by the detection equipment. 

Injection Operation 

The objective of the Rock Springs test project was to place 20,000 cubic 
yards of sand in underground voids through a single injection hole. This 
should constitute a convincing demonstration inasmuch as the quantity of fill 
that could be injected under the existing method of blind flushing was esti- 
mated at an average of 100 cubic yards per hole. 

It was estimated that 20,000 cubic yards of sand would fill mine voids 
within an average radius of 210 feet from the injection hole, based on an 
average 6-foot thickness of the coalbed and 65-percent extraction of the coal. 
The corresponding surface area overlying the mine openings to be backfilled 
was calculated to be 3.2 acres. 

The material used for backfill was fine sand of wind-blown origin avail- 
able from a nearby deposit. The sand was screened at the pit to reject parti- 
cles larger than one- fourth inch in diameter, and pieces of debris. The sand 
was transported by truck to the injection site and stockpiled. 

Water to form the slurry was obtained from the flooded mine by means of 
two wells located about 325 feet downdip from the injection well. The mine 
water contained about 13,500 ppm dissolved solids but had no disagreeable 
odor and was not highly corrosive (3). Two submersible water pumps, each with 




FIGURE 8. - Photograph of the plant site at the first test of the pumped-slurry mine backfilling 
technique. 



17 



capacity of 4,000 gpm, pumped water to the mixing tank. A reserve supply of 
water for purging the system was maintained in four storage tanks adjacent to 
the mixing tank. Figure 8 is a view of the plant site, and figure 9 shows the 
equipment installation in relation to streets. 

Water entered the mixing tank at an average rate of 5,500 gpm and mixed 
with the sand, which was supplied at the rate of 120 cubic yards per hour, to 
form a slurry. A slurry pump impelled the slurry through the 13-3/8-inch-ID 
injection pipeline at an average velocity of 17 fps. The injection borehole 
was cased with 13-3/8-inch-ID pipe to within 5 feet of the mine roof, which 
was 116 feet below the surface. 



X 



6' chain link fence 



co 

in 



o 
2 



13 3/8 slurry line 



Material hopper 
and conveyor ■> 

J- 7 




if 



# 



/ Slurry mixing tanks 



FIGURE 9. - Plan view of installed equipment at the plant site and adjoining streets. 



18 



The sand slurry was successfully injected into the mine workings over a 
10-day period. Operations were scheduled for 24 hours per day, but actual 
daily injection time ranged up to a maximum of 21 hours because of electrical 
and mechanical problems. The solids concentration in the slurry averaged 
0.8 pound per gallon of water. No resistance to injection was encountered, 
and the pressure measurements made at the top of the injection borehole were 
below atmospheric (vacuum) throughout most of the period. 

During the last 12 hours of injection, it was planned to increase the 
sand concentration and decrease the discharge rate in order to achieve com- 
plete filling up to the mine roof. When the flow rate decreased to about 
3,000 gpm, however, the sand in the mixing tank was not kept in suspension 
and the discharge pump became plugged after 19,500 cubic yards of sand had 
been injected. 

Evaluation 

The demonstration at, Rock Springs proved that the pumped-slurry injection 
process could successfully emplace approximately 20,000 cubic yards of sand in 
mine voids from a single injection hole. In fact, there seemed little doubt 
that more sand could have been emplaced had the initial injection rate been 
maintained. 

Much more difficult to evaluate was the extent (both laterally and 
vertically) to which the open spaces in the mine had been filled. Prior to 
the injection operation, four holes had been drilled and cased to be used for 
observing the results of the filling process. Difficulties were experienced 
in determining depths to sand fill in these holes, however, because the sam- 
pler could not penetrate the sand without churning it into a "quick" condition. 
Interpretations were further complicated by the probability that the open 
voids encountered in these drill holes were actually caved spaces above the 
mine roof rather than rooms at mine level. 

After completion of the injection project, the Bureau of Mines instituted 
an evaluation program based on the drilling of 36 additional holes. In fig- 
ure 10, the heavy line encloses the distribution of fill as determined by the 
evaluation team on the basis of all available subsurface data. The circle 
represents the area with a radius of 210 feet from the injection hole that 
was the planned area of fill distribution. The actual area backfilled in the 
mine was 2.8 acres as compared with the predetermined area of 3.2 acres. 
Restriction of fill to the west is believed to have been caused by human- 
engineered obstructions (air stoppings) for ventilation control that were 
not shown on the mine map. 

Evidence as to vertical completeness of the fill was the most difficult 
to collect. In the two holes closest to the injection hole, at distances of 
about 60 and 100 feet, mine voids were completely filled. In other holes, the 
degree of vertical filling was hard to interpret, in part because the original 
mine roof had caved, extending the opening upward from 7 to 10 feet, and the 
caved material formed mounds on the original mine floor. In some drill holes, 
sand was found interbedded with caved rubble; in others, the sand was in the 



19 




Scale, feet 

FIGURE 10. - Map of mine workings at the site of the first 
demonstration project in Rock Springs, Wyo., 
showing pattern of sand deposition in the 
mine voids. Dashed circle shows area of 
planned backfill. 



caved space above the rubble. 
In a few holes, sand was blown 
up the hole when the drill 
encountered a cavity, indi- 
cating air pockets trapped 
at the top of the caved 
spaces. In a hydraulic fill- 
ing operation, such air- 
filled voids would remain 
unfilled. 

The relative density of 
the fill was found to decrease 
from the injection point. 
Tests by the Wyoming Highway 
Department indicated average 
in-place density of about 
127 pounds per cubic foot 
with relative density values 
ranging from 36 to 81 per- 
cent. This range included 
mixtures of sand and shale 
rubble (14). 

The degree of subsidence 
control effected by this 
project is difficult to eval- 
uate. The emplaced fill in 
such a relatively small area 
of the mine would not be 
expected to completely pre- 
vent subsidence throughout 
the area because of apparent 
decreasing height and density 
of the fill away from the 
injection hole. Any further 
subsidence that might take 
place in peripheral areas, 
however, would tend to be 
somewhat less than in com- 
parable areas that had not 
been backfilled. The pro- 
cess was considered to hold 
sufficient promise to jus- 
tify further experimentation. 

Cost 

The cost of the demon- 
stration project, including 



20 



all the extras associated with a first-time application, was $9.00 per cubic 
yard ($6.75 per ton) broken down as follows: 

Project planning $0. 158 

Investigation of mine openings . 542 

Preparation of wells and manifolds 2.350 

Fill material and handling 2. 832 

Injection 2. 818 

Site restoration and reporting .300 

Total cost per cubic yard 9. 000 

These costs reflect the demonstration of the process only and do not 
include the preceding site studies or the subsequent evaluation program. The 
cost per acre, about $62,500, is excessive, of course, and reflects the disad- 
vantage of applying the method wherever only a relatively few mine voids are 
to be backfilled. The success of the test, however, which proved that 
19,500 cubic yards of sand could be inserted into mine voids through a single 
borehole, promised invaluable aid to mining communities, fully justifying the 
cost of the experiment. " 

Reports 

The Department of Housing and Urban Development, in cooperation with the 
city of Rock Springs, Wyo. , arranged with Candeub, Fleissig and Associates, 
Consultants, Newark, N. J. , to make a study of the demonstration project and 
a comprehensive report on the backfilling technique and its possible applica- 
tion in other mining areas of the Nation (_3) . . 

After the demonstration, the city of Rock Springs initiated a broad study 
of its overall community needs under the Community Renewal Program with assist- 
ance from the Department of Housing and Urban Development. This study has led 
to the establishment of priorities not only for subsidence control, but also 
for the treatment and renewal of other sources of deterioration and blight (19) . 

A comprehensive study of the general geology and underground mining in 
the Rock Springs area, in conjunction with the Community Renewal Program for 
the city, was made for the purpose of determining the economic feasibility of 
backfilling the mined-out areas. The report was made by the firm of Johnson- 
Fermelia & Crank, Inc., Consulting Engineers and Land Surveyors, Rock Springs, 
Wyo., and was completed March 1, 1972 (9) . Two recommendations are included 
in the report: (1) That an extensive exploratory drilling program be con- 
ducted in various parts of Rock Springs to verify existing mine map data; and 
(2) that Federal funding be obtained for a large-scale mine backfilling proj- 
ect under the Belmont-Kerback area to assess the effectiveness of the process 
in arresting the continuing subsidence in that area. 

FIRST LARGE-SCALE TEST 

Scranton, Pa. , was selected by the Bureau as the locale for the full- 
scale pumped-slurry demonstration project in 1972-73 because of its subsidence 
history and the active local interest in subsidence control. Population 



21 



centers in the Anthracite region of northeastern Pennsylvania have had a his- 
tory of subsidence problems as a result of multiple-bed mining over a period 
of 150 years. Scranton is the largest of the cities in the Anthracite region. 
As many as 11 different coalbeds have been mined under Scranton, and most of 
the central part of the city overlies 6 mined beds. The demonstration project 
site is in an area of potential subsidence due to past mining in five aban- 
doned coalbeds. Subsidence had not yet become apparent at the ground surface 
in this area, and caving below ground was not believed to be sufficient to 
block effective movement of slurry. 

A 30-acre residential area was stabilized by injecting about 450,000 cubic 
yards of crushed mine refuse into two coalbeds through five injection bore- 
holes. Nearly 200,000 cubic yards was injected through one borehole from 
which the material moved into the mine workings on all sides; the injected 
material reached a maximum lateral distance of 640 feet and filled mine open- 
ings from floor to roof. The new method, designed for inundated mines, proved 
successful also in mine workings above mine-water pool level (22). 

HYDRAULIC MODEL STUDIES FOR BACKFILLING MINE CAVITIES 

In October 1973, the Bureau of Mines arranged for model studies of the 
pumped-slurry method of backfilling mine cavities to be conducted by the 
U.S. Bureau of Reclamation at their hydraulic laboratory in Denver, Colo. The 
purpose of the studies was to obtain qualitative and quantitative data on the 
deposit pattern of fine sand, such as that at the Rock Springs projects, when 
used for backfill material and injected into cavities of a simulated coal mine. 
The results of the first series of 18 tests are given in a report entitled, 
"Hydraulic Model Studies for Backfilling Mine Cavities," by E. J. Carlson of 
the U.S. Bureau of Reclamation, which is included as an appendix in Bureau of 
Mines Information Circular 8667 (22) . 

A second series of model studies of the pumped-slurry method using the 
same fill material and simulating conditions not covered in the first series 
was made by the Bureau of Reclamation for the Bureau of Mines in 1974. That 
report was also by E. J. Carlson, U.S. Bureau of Reclamation; dated March 1975, 
it is included as an appendix to this report. 

THREE LARGE-SCALE PROJECTS AT ROCK SPRINGS 

The success achieved with the pumped-slurry technique at Rock Springs and 
at Scranton, where for the first time extensive mine voids in two beds under 
about 30 acres were filled from only five injection boreholes, prompted the 
Bureau to initiate several more large-scale subsidence control demonstration 
projects in both the Anthracite region and Rock Springs. Experiments were 
needed to determine the various types and sizes of material as well as the 
optimum range of conditions under which the pumped-slurry method could be 
used more effectively in minimizing adverse environmental problems associated 
with underground coal mining. In addition to the original 20,000-cubic-yard 
test, this report covers the three large-scale demonstration projects that 
were conducted in Rock Springs during the period 1973-76. 



22 



It became apparent, however, that the Bureau's proposed subsidence con- 
trol efforts at Rock Springs would proceed in an atmosphere of unprecedented 
industrial expansion. Areas near the city saw the construction of a $500 mil- 
lion coal-fired steam-electric powerplant; development and expansion of trona 
operations by four chemical companies; oil and gas exploration; and stepped-up 
activities relating to the extraction of coal and uranium. The population of 
Rock Springs rose from a more or less stable 12,000 in 1970 to 26,000 in 
January 1974. In addition to subsidence problems, therefore, the city was 
obliged to cope with crowded schools, traffic jams, lack of housing, rising 
crime rates, overloaded public facilities, and insufficient tax revenues. 

Under the authority of the Organic Act, the Bureau of Mines entered into 
a cooperative agreement with the city of Rock Springs to conduct large-scale 
demonstrations of the pumped-slurry method that would also provide surface 
support for the subsidence-prone critical areas of the city. The agreement 
provided that the Bureau design, direct, and supply the funds to accomplish 
the work by contracts obtained through Federal rules and regulations governing 
competitive procurements. The Bureau, through its contractors, was required 
to conduct all phases of the work with a minimum of inconvenience to residents 
and with as little adverse effect upon the environment as possible. The city 
was required to grant the Bureau, its officials, contractors, or employees, 
the right to enter streets, roads, and any other land owned or controlled by 
the city overlying or adjacent to the project areas for the purpose of con- 
ducting the injection operations. The city also furnished without charge all 
licenses, permits, or other authorizations required in the conduct of project 
work, which included detours for local traffic and police surveillance of 
Bureau property. City engineers and utility personnel also provided accurate 
information regarding locations of utility pipelines, wires, cables, and sewer 
lines together with maps showing ownership of property within and adjacent to 
the project areas. 

Drilling Operations 

All drilling and reaming of boreholes that were used for water wells, 
slurry injection, exploration, and monitoring was done with rotary drills 
equipped with efficient dust control mechanisms. While drilling was done with 
air, insofar as possible, water (mist) injection was also used to reduce the 
amount of dust discharged into the atmosphere. After drilling through uncon- 
solidated overburden, all boreholes that were used for water supply and injec- 
tion were cased and cemented to a point 4 feet into the rock stratum. From 
here drilling was continued at sufficient diameter to accommodate the required 
casing to the mine opening. Such boreholes were collared, sealed to the 
standpipe, and fitted with removable pressure-tight caps at the surface. For 
the monitoring boreholes, casing was not installed below the point 4 feet into 
the rock stratum. Logs of all the boreholes, and particularly the depth at 
which the bit broke through to the mine opening as well as the height of the 
mine opening, were furnished to the Bureau of Mines. Figures 11 and 12 are 
typical cross sections of the various boreholes used in the projects. 

Boreholes that penetrated into solid coal and therefore not usable were 
filled with drill cuttings to within 3 feet of the top of solid rock, then 



23 



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24 



with 3 feet of concrete. From this point the hole was filled with cuttings or 
soil within 1 foot of the surface. The remaining 12 inches was filled with 
dirt, asphalt, or concrete depending upon the kind of material originally pres- 
ent. After the injection and monitoring boreholes served their purpose, they 
were filled with sand to within 4 feet of the top of rock, then with 4 feet of 
concrete to the top of rock. From this level the hole was filled with sand to 
the surface, except in those instances where the boreholes are in roadway sur- 
faces, in which case they were filled with concrete or the type of roadway 
material present for at least the top 12 inches. Where not used to backfill 
boreholes, the mud and drill cuttings produced during drilling were removed 
from the various sites. Wherever possible, casings were removed after the 
work was completed. 

In addition to providing usable boreholes, the purpose of the exploratory 
borehole drilling was to outline and define, to the extent necessary, the size, 
location, and condition of the abandoned coal mines underlying the proposed 
project areas. Because of apparent errors in maps of the coal workings and 
the inaccessibility of the mines for inspection or resurveying, the purpose 
of the proposed drill holes was also to help establish the position of mine 
workings with relation to the surface. In a few parts of the project areas 
the locations of mine voids were represented accurately; however, in most of 
the project areas, after the drill holes encountered pillars, it became neces- 
sary to reposition the mine maps with respect to surface features until the 
drill holes penetrated voids on a more frequent basis. A device of this kind 
may be applied successfully in other mining areas where similar inaccuracies 
exist. 

Slurry Components 

Fine sand from various deposits of wind-blown origin in the vicinity of 
Rock Springs was used as fill material in all of the demonstration projects at 
Rock Springs. Except for screening out rocks and other debris, little else 
was done in preparing the sand for injection into mine voids. The grain-size 
gradation of a typical sample of sand used in the various projects and in the 
model studies is shown graphically in appendix figure 3. Data on settlement 
tests, in-place density, and bearing capacity of the fine, uniform, blow sand 
(median size, ~ 0.14 mm) from the Rock Springs area are also given in the 
appendix. 

Specifications regarding the use and reclamation of borrow areas where 
sand was obtained were established in accordance with requirements of the 
Bureau of Land Management, U.S. Department of the Interior, and incorporated 
in the contracts for the three large-scale demonstration projects. Sandfill 
for Project 1 (Dow Chemical Co.) and for Project 3 (M. J. Bober Co.) was taken 
from deposits controlled by the Bureau of Land Management. Sand for Project 2 
(WHAN Engineering and Construction, Inc.) was obtained from a deposit on pri- 
vately owned land of the Upland Industries Corp., Rock Springs, Wyo. In all 
instances, the following stipulations regarding the use of borrow pits applied: 



25 



1. Topsoil, if present, will be removed and stockpiled. 

2. At such time when materials (blow sand) has been removed and rehabil- 
itation is underway, the floor of the pit shall be leveled to coincide with 
the existing watershed slope. If no other use is made of the screening 
rejects they will be scattered over the pit floor. Topsoil will be replaced. 
Pit slopes will be 3:1 ratio. Haulage roads will be leveled and berms 
smoothed. Haulage roads and materials pit will be seeded with crested wheat 
at 5 lb per acre and thickspike wheatgrass at 3 lb per acre. Seeding will be 
accomplished by using a grass seed or grain drill with grass seed attachment. 
Seeding shall be at 1/2-inch depth and will continue until a satisfactory 
stand is established and approved by the Bureau's representative. 

3. Seeding will be done in the fall after September 15 and before 
freeze-up. 

4. Leveling, grading, and seeding shall be completed within 1 year 
after the completion or termination of any operation hereunder. 

5. Road crossings shall not impede drainage; either culverts should be 
installed or approaches cut down to drainage bottom level. 

6. In the event an archeological site is unearthed, the Bureau of Mines 
shall be notified immediately. 

7. Signs shall be properly posted to warn the public of the extraction 
area. 

8. Proper precaution will be taken at all times to prevent and suppress 
fires. The contractor will be held responsible for suppression costs for the 
fires caused through negligence of his employees or subcontractors. 

9. Should the material be slurried at the site, water shall be 
controlled. 

10. Upon termination of the pit, the contractor shall submit a final 
accounting of the total material removed. 

The source of water, which is needed in large quantities for the pumped- 
s lurry injection technique, was the local mine-water pool beneath the test 
site and beneath the three large-scale projects. Advantages of using mine 
water include its availability, low cost, and the probability that withdrawal 
and injection of large quantities of water from, and into, the same body of 
water would create minimum disturbance of the subsurface equilibrium. The 
mine water contained about 13,500 parts per million (ppm) dissolved solids 
but had no disagreeable odor and was not highly corrosive (9) . 

Probe boreholes that intersected the mine-water pool at required pump 
drawdown depths at or near the various project sites were enlarged to accom- 
modate large casings in which deep-well or submersible pumps were installed. 
Specifications for the first and second large-scale projects required delivery 



26 



of a minimum of 4,000 gallons per minute (gpm) to the mixing tank. This was 
increased to a minimum of 5,000 gpm for the third large-scale project to allow 
for an increased volume of slurry injection. Data on solids, water, and slurry 
covering the range of quantities utilized at the various projects are given 
in table 1. 



TABLE 1. - Data on solids, water, and slurry 



Slurry water to 
solids ratios 



4,000 gpm 

(16.66 
tons/min) 



5,000 gpm 

(20.82 
tons/min) 



Water pumped 



6,000 gpm 

(24.99 
tons/min) 



7,000 gpm 

(29.15 
tons /min) 



8,000 gpm 

(33.32 
tons/min) 





SOLIDS INJECTED: 200 TONS/HR, 3.33 TONS/MIN, 


307 GPM 






5.0 
13.0 


6.3 
16.3 


7.5 
19.5 


8.7 
22.8 


10.0 
26.0 





SOLIDS INJECTED: 350 TONS/HR, 5.83 TONS/MIN, 


538 GPM 






2.9 

7.4 


3.6 
9.3 


4.3 
11.2 


5.0 
13.0 


5.7 
14.9 





SOLIDS INJECTED: 450 TONS/HR, 7.50 TONS/MIN, 


692 GPM 






2.2 
5.8 


2.8 
7.2 


3.3 
8.7 


3.9 
10.1 


4.4 
11.6 



Sand: 



Bulk density 100 pcf (loose) 38.3 pet voids; 61.7 pet solids 
2.6 (specific gravity) X 62.4 lb = 162.2 pcf (solid) 
162.2 t- 7.48 gal/ft 3 = 21.68 lb/gal (solid) 



Water: 



4,000 gpm X 8.33 lb/gal 
5,000 gpm X 8.33 lb/gal 
6,000 gpm X 8.33 lb/gal 
7,000 gpm X 8.33 lb/gal 
8,000 gpm X 8.33 lb/gal 



33,320 lb /min 
41,650 lb /min 
49,980 lb /min 
58,310 lb/min 
66,640 lb/min 



16.66 tons/min 
20.82 tons/min 
24.99 tons/min 
29.15 tons/min 
33.32 tons/min 



Injection rate of solids : 
200 tons/hr = 3.33 tons/min 
350 tons/hr = 5.83 tons/min 
450 tons/hr = 7.50 tons/min 



307 gal /min 
538 gal /min 
692 gal /min 



Injection rate of slurry by volume (two examples) : 

Water + Solids = Slurry 

5,000 gpm 3.33 tons/min 5,307 gpm 
7,000 gpm 5.83 tons/min 7,538 gpm 

Mixing Plants and Slurry Pumps 

Equipment at a typical mixing plant site at Rock Springs included a hop- 
per, variable-rate feeder, belt weighing scale, mixing tank, diesel-powered 
slurry pump, piping and valves, and other auxiliary apparatus necessary to 
operate, control, maintain, and monitor the backfilling operation. Each 



27 



mixing tank was constructed and arranged in such a manner that it was capable 
of receiving measured and weighed sand at a rate of at least 1,600 tons per 
shift and mine water ranging from 3,000 to 6,000 gpm from electric-powered 
deep-well pumps. Mixing tanks were fitted with manifolds connecting the 
incoming water supply with a series of nozzles, producing sufficient turbu- 
lence to maintain the solids in suspension. In each tank the suction line 
was fitted with a snorepiece that picked up the suspended material (slurry) 
and conveyed it to the impellers in the slurry pumps. From here the slurry 
was pumped through distribution lines to the various injection boreholes. 

In each case all operations involved in the mixing, pumping, and injec- 
tion processes were controlled from a conveniently located instrument panel. 
This included the rate of waterflow, the rate at which sand was introduced, 
the amount of slurry discharged, cumulative weight of solids injected, and 
the pressure both at the slurry pump and injection borehole. 

Because of the experimental nature of the work and varying conditions to 
be encountered at the three projects, the specifications allowed a certain 
amount of flexibility in conducting the work. Each contractor and/or his 
engineer was encouraged to select and operate the equipment so as to discover 
the optimum range of such variables as volume, velocity, and density of slurry 
flow in relation to changes in pump head, pressure at the injection borehole, 
length of discharge line, and depth to mine voids. 

PROJECT 1 

Subsequent to the completion of the first successful large-scale applica- 
tion of the pumped-slurry mine backfilling process at Scranton, Pa. , Congress 
in fiscal year 1973 appropriated $700,000 for the Bureau to conduct a similar 
large-scale demonstration at Rock Springs, Wyo. 

In designing the project, the Bureau selected critical areas 1, 7, and 8 
because the estimated cost of the work and material to fill the mine voids 
would approximate available funds. Moreover, it was hoped that providing sup- 
port in the mine workings under area 1 might arrest the continuing subsidence 
occurrences in and near the Connecticut Avenue-D Street section. The three 
critical areas were located in the southeast portion of Rock Springs above 
both flooded and dry mine voids in the No. 7 seam of the abandoned No. 2 Mine, 
Central Coal and Coke Co. The old mine workings were believed to be under 
less than 300 feet of overburden, and this was to be the first time that sand 
fill would be used in the pumped-slurry process to fill both inundated and 
dry coal mine voids. It was proposed to demonstrate that approximately 
150,000 cubic yards (200,000 tons) of screened sand could be introduced 
into the designated mine voids under critical areas 1, 7, and 8, a total 
area of about 33 acres (see fig. 13). 

Prior to preparing specifications for the backfilling project, explora- 
tory drilling was conducted in the three critical areas. It was necessary to 
know the nature and depth of overburden; condition of the mine workings; and 
whether or not the available maps of the mine workings show a true relation- 
ship with overlying surface features. The Bureau, therefore, through Federal 



28 




300 600 900 

_l I I 



Scale, feet 

FIGURE 13. - Map of a portion of Rock Springs showing locations of 
critical areas 1, 7, and 8 in relation to mine workings 
of the No. 7 Seam. 



29 



Government competitive contracting and bidding procedures, awarded a drilling 
contract on January 26, 1973, to the Delta Drilling and Development Co. , Dana, 
Ind. , in the amount of $39,636.00. Forty usable boreholes, large enough to 
accommodate 4-inch-ID plastic casing, were eventually drilled in the project 
areas to provide the necessary information. Depths to the mine voids ranged 
from 85 to 272 feet, and alluvium ranged between 10 and 45 feet. Actual cost 
of the work was $37,707.02. Under the ensuing backfilling contract, eight of 
the boreholes were reamed to larger sizes to provide water-supply wells and 
injection boreholes. The remaining 32 boreholes were used to monitor the 
backfilling progress. 

As in the earlier test project, it was believed that while the explora- 
tory drilling provided approximate vertical depth of the void at each borehole, 
the lateral extent of the voids was not apparent. The Bureau, therefore, 
employed the Dowell Division of the Dow Chemical Co. to explore the extent of 
mine passageways between boreholes with "Sonar Caliper" surveys. Eleven bore- 
holes were probed with the instrument, which confirmed to some extent the con- 
tinuity of the voids. The cost of the work was $5,135.15. 

Through competitive contracting procedures, the Bureau, on June 4, 1973, 
awarded a pumped-slurry backfilling contract to the Dowell Division of the Dow 
Chemical Co. in the amount of $721,777.50 to provide surface support under 
areas 1, 7, and 8. The contract was modified on June 29, 1973, to allow for 
the drilling, casing, and cementing of a necessary alternate injection bore- 
hole, increasing the total contract amount by $7,924.00 to a new total of 
$729,701.50. 

During the mobilization and installation of equipment, the contractor 
elected to locate the injection plant at the borrow site rather than use 
trucks to haul the sand about 1-1/2 miles to the Bureau's suggested plant 
site location. This of course involved pumping the slurry directly from the 
borrow site through a network of pipelines an average distance of 2.1 miles 
to the injection boreholes. Other work included under mobilization involved 
preparation of the borrow pit, installation of the water-supply system, erec- 
tion of the injection plant facilities, and construction of the 14-inch-ID 
slurry pipeline. At this time, the injection boreholes were also provided 
and suitably capped to await being used. The slurry injection was started 
August 24, 1973. 

Two submersible water pumps were installed, one in each of the two water 
wells just north of critical area 1. The pumps were 137 feet below the sur- 
face and about 35 feet below the average mine-water pool level. The pumps 
were rated at 4,000 gpm each, at a dynamic head of 150 feet, and were driven 
by 200-hp electric motors. The water was pumped from the pool in the No. 7 
Seam and impelled with the aid of a 1, 200-hp, diesel-powered booster pump 
through 1-1/2 miles of 14-inch-ID steel pipeline to the mixing plant at the 
borrow site. Because of the friction head in the relatively long pipeline, 
only an average of 3,600 gpm was delivered to the mixing plant. 

The sandfill material was obtained from a nearby deposit on Federal land 
controlled by the Bureau of Land Management (BLM). This designated site was 



30 



made available to the contractor through a "Free Use Permit* 1 between the BLM 
and the Bureau of Mines. Removal of the fill material and subsequent reclama- 
tion of the borrow site was conducted in accordance with the stipulations 
required by the BLM, which were made part of the demonstration project speci- 
fications by an amendment. Adequate screening equipment was installed to 
reject rock particles larger than one-half inch as well as other foreign 
debris. 

At the borrow pit sand was loaded by bulldozers and scrapers into a hop- 
per and fed through a reciprocating plate onto a conveyor, then weighed as it 
was being delivered to the steel mixing tank. Water from the submersible 
pumps entered the tank through a manifold and a series of nozzles producing 
sufficient turbulence to maintain the solids in suspension. The slurry pump, 
of high-chrome cast iron, was a 1,200-hp, diesel-powered, centrifugal dredge 
service pump, capable of pumping slurry at a rate of 8,400 gpm at a head of 
175 feet. Figure 14 is a view of the slurry injection plant at the site of 
the borrow pit. Although the slurry was transported slightly down-grade, the 
friction head that developed in the long pipelines reduced the volume of 
slurry delivered to the injection holes to an average of about 4,000 gpm. 

The slurry distribution pipelines were buried beneath the busier streets 
and intersections to minimize the impact on traffic. On less-traveled streets 
they were laid along curbs where arrangements with property owners permitted 
construction of ramps over pipelines at entrances to driveways. The locations 
of the water-supply wells, the water-supply pipeline complete with booster 
pump, the borrow pit, injection plant, housing for recording and control equip- 
ment, slurry distribution pipelines, and injection boreholes are shown in 
figure 15. 




FIGURE 14. - View of the slurry mixing plant installation at the site of the borrow pit about 
2 miles from the built-up areas of the city. Note natural rock outcropping in left 
background off project limits and slurry pipeline in left foreground. 



31 




FIGURE 15. - Map showing location of water-supply wells and pipeline, slurry pipelines, and 
injection boreholes in relation to critical areas 1, 7, and 8. (The distance to 
the injection plant at borrow pit is indicated.) 



32 



A total of 152,1*61 tons of solids was injected into the three separate 
areas of the No. 7 mined coalbed. The total surface area overlying the back- 
filled portions of the coalbed was estimated to be approximately 33.2 acres. 
Seven injection boreholes were used in the backfilling. Injection bore- 
holes D-33 and D-35 were located in inundated mined areas and boreholes D-15, 
D-l, D-28, D-90, and D-113 in dry areas. A total of 62,119 tons of fill mate- 
rial was injected into inundated mine voids, and a total of 90,348 tons was 
injected into dry voids. The progress of backfilling is summarized in table 2, 

TABLE 2. - Injection data on the initial large-scale demonstration 

project, 1973-74 



Project area 
and acreage 



Injection 
borehole 



Approximate injection periods 



Total 

injections, 

tons 



1, 21.10 acres 



D-33. 
D-15. 
D-35. 
D-l.. 
D-28. 



7, 3.90 acres 



D-113, 



Aug. 24-Sept. 10, Sept. 17-19, 1973. 

Sept. 10-17, 22-24, 1973 

Sept. 19-22, 25-26, 1973 

Sept. 24-25, 1973 

Sept. 25, 1973 



Oct. 3-13 and Nov. 11-12, 1973. 



8, 8.30 acres j D-90 1 Nov. 12-18, 1973, and Apr. 2-28, 1974. 



54,726 

15,914 

7,393 

1,105 

1,579 



26,054 



45,696 



The slurry operation began on August 24, 1973, and was completed on 
April 28, 1974. Two major interruptions in the operation were experienced: 

1. Breakdown and subsequent overhaul of the diesel engines operating the 
water-supply booster pump and slurry pump, resulting in downtime October 14 
through November 11, 1973. 

2. The shutdown of operations in anticipation of severe winter condi- 
tions that would cause the freezing of pipelines, valves, etc. , from Novem- 
ber 18, 1973, through April 1, 1974. 

Backfill material was distributed in the mine workings throughout the 
three critical areas, comprising about 33.2 acres, as indicated by observa- 
tions of the 32 monitoring boreholes. The presence of fill material in moni- 
tor holes at or above mine-roof levels indicated that filling of the mine 
voids within the project areas was essentially complete from floor to roof. 
The estimated extent of the backfilling in the three areas is consistent with 
the computed volume of void space in the mine workings and with the quantity 
of fill material that was injected. Undoubtedly some fill material extends 
beyond the limits of the completely filled voids where it would be expected 
to conform to the submerged angle of repose of the sand, about 30°. 



In the three project areas, the dip of the No. 7 Seam ranged up to 5°. 
The monitoring results, which were later confirmed by laboratory model studies 
(see appendix), indicated that in inundated voids the water serves as a flow 
medium yielding a balanced updip-downdip distribution pattern. When it became 



33 



apparent that injected slurry in dry mine voids traveled more readily downdip, 
subsequent injection boreholes in dry areas were placed at higher seam alti- 
tudes to take advantage of the downdip flow tendency. Data on the behavior 
of slurry deposition in a dipping bed in both flooded and dry mine cavities 
are described in more detail in the appendix. 

Only two significant operational design problems were encountered during 
the demonstration. One was the detection of slurry through the water supply 
wells. This indicated that the mine water was in fact returning to the mine 
pool as planned, and that recirculation of the mine water was occurring. 
Because excessive slurry recirculation through the water supply wells could 
cause damage to well pumps, it was concluded that in future demonstration 
projects the water supply wells should be located in a more remote area so 
as to prevent any recirculation of slurry. 

On one occasion some entrained air forced slurry to the surface through 
a plastic-cased monitoring borehole. No property damage was reported; however, 
slurry was discharged into the street and required a cleanup. It was con- 
cluded that henceforth all monitoring boreholes would be cased with steel pipe 
to a point 4 feet into the top of rock and that accessible steel caps would be 
welded on to provide for a pressure-tight covering. 

The cost of the hydraulic backfill demonstration project, in which 
152,46:7 tons of sandfill material was injected, was $729,000. Reductions in 
the unit cost of the pumped-slurry technique were anticipated as additional 
experience was gained in applying the method. Total cost of the project, 
including exploratory drilling and sonar surveys, was $772,543.67 — a unit 
cost of $5.07 per ton. 

The demobilization of the operation, including cleanup and restoration of 
all project work areas, was conducted in accordance with the contract specifi- 
cations. Restoration of the borrow pit was completed in accordance with the 
stipulations set forth by the Bureau of Land Management and included seeding 
of the borrow area in fall 1974 and a verification inspection of the satisfac- 
tory growth after one complete growing season in spring 1976. 

PROJECT 2 

In fiscal year 1974 Congress provided $700,000 for the Bureau of Mines 
to continue utilizing the pumped-slurry backfilling technique in areas of Rock 
Springs where subsidence control would be beneficial to the city. 

The demonstration project site was selected in an area of potential 
subsidence due to extensive mining in No. 1 Seam of the abandoned No. 1 Mine, 
Union Pacific Coal Co. This would be the first time that the pumped-slurry 
method of backfilling mine voids would be demonstrated in a central downtown- 
business area of a city. The objective of the project was to fill both 
flooded and dry coal mine voids with solids that would support overburden and 
pillars so as to minimize possible future damage to surface structures, 
streets, and public facilities. The voids to be backfilled by this project 
consisted of critical area 14 and about one-half of critical area 11. These 
areas are shown in figure 16. 



34 




FIGURE 16. - Mapofa portion of Rock Springs showing location of critical area 14 and a part 
of critical area 11 in relation to underground mine workings in the No. 1 Seam. 



35 



To accomplish this objective, it was proposed to demonstrate that at 
least 150,000 tons of screened sand could be introduced into the flooded 
and/or otherwise inaccessible mine voids beneath the areas described above, 
a total planned area of about 28 acres. The project work was conducted in 
two phases: Phase I--Exploratory Borehole Drilling; Phase II — Hydraulic Back- 
filling. Execution of Phase II was contingent to a degree upon a successful 
Bureau evaluation of the results of the work conducted under Phase I. 

The purpose of the exploratory borehole drilling was to outline and 
define, to the extent necessary, the size, location, and condition of the 
abandoned coal mines underlying the proposed project areas. Because of 
apparent errors in maps of the coal workings and the inaccessibility of the 
mines for inspection or resurveying, the proposed drill holes would help to 
establish the position of mine workings with relation to the surface. 

Through Federal Government competitive contracting and bidding procedures, 
the contract to conduct both phases of the proposed work was awarded April 4, 
1974, to the WHAN Engineering and .Construction, Inc. , Bismarck, N. Dak. , the 
lowest bidder, in the amount of $666,667.00. Provisions were included in the 
contract to create a minimum of disturbance to the boom-town environment exist- 
ing in downtown Rock Springs. 

Phase I involved the drilling of exploratory boreholes throughout the 
proposed project area. Twenty-eight holes encountered mine openings, and 
although some caving had been noted, it was not believed to be sufficient to 
block effective movement of slurry. Depths to the mine voids ranged from 
about 30 to 160 feet, and alluvium depths ranged between 15 and 60 feet. Some 
of the holes intersected the mine-water pool, which helped to determine the 
location and depth of the pool in the project area. This information, in con- 
junction with a study of the mine map, gave direction on the best probable 
locations for water-supply wells, injection boreholes, and monitor holes. It 
was apparent also from the drilling that the remaining pillars in the project 
area were somewhat smaller than had been expected. The void space to be 
filled, therefore, was recalculated adding 15,000 tons of fill material to 
the previously estimated quantity for a new total of 165,000 tons. To accom- 
modate handling the increased quantity, the contract was modified June 28, 
1974, to increase the total contract amount to $680,167.00. An attempt to 
substantiate exploratory drilling findings of mine conditions by sonar surveys 
as in Project No. 1 was not deemed necessary in Project No. 2. 

Under Phase II, two of the exploratory boreholes completed under Phase I 
were reamed to a larger size and converted to water-supply wells for the back- 
filling process. Twenty boreholes were cased with 4-inch-ID steel casing pipe 
and capped with removable pressure-tight covers to be used as monitoring 
boreholes. 

Phase II also involved the installation of deep-well pumps, 1,300 feet of 
water-supply pipeline, the slurry injection plant, and some of the slurry dis- 
tribution pipelines at sites designated by the Bureau of Mines on the right of 
way of the Union Pacific Railroad. At the request of the Bureau, the railroad 
company granted permission of the use of its property for water-supply wells 



36 




FIGURE 17. - Map of the project area showing location of the slurry plant in relation to 
water-supply wells and pipeline, slurry pipelines, and injection boreholes. 



37 



and much of the installation of the hydraulic backfilling facilities. The 
right of way was conveniently located to the project areas and provided ade- 
quate space for the operations, including room for a sand stockpile, such that 
the environmental impact upon the community was minimal. Use of the property 
for these purposes was formally arranged between the contractor and the Union 
Pacific Railroad. Slurrying and injections of solids into the mine voids began 
August 7, 1974. The location of the plant, pipelines, and injection boreholes 
are shown in figure 17. 

Two deep-well pumps lifted the water from the mine at a minimum rate of 
4,000 gpm and impelled it through about 1,300 feet of 14-inch-ID welded steel 
pipe to the site of the mixing plant. Two deep-well turbine pumps were used, 
one in each well, at points located 95 feet below the surface and 35 feet 
below the average mine-water pool level. The pumps were rated 4,000 gpm each 
at a dynamic head of 150 feet and driven by 200-hp electric motors. A part of 
the 1,300 feet of pipeline from the deep-well pumps to the mixing plant 
included crossing under the mainline of the Union Pacific Railroad. A 120-foot, 
18-inch-diameter bore was used for this purpose with the permission of the 
railroad company. 

Sand was obtained from a borrow pit on property of the Upland Industries 
Corp. , Rock Springs, Wyo. , and paid for by the Bureau at a rate of $0.10 per 
ton. Figure 18 is a view of equipment being operated at the borrow pit. The 
sand was screened to remove rock particles and other debris in excess of 
5/8-inch size and loaded into 30-ton bottom-dump trucks, which hauled the 
sand 2.5 miles to the site of the mixing plant. The change to a larger size 
screen from that used in the first large-scale project reflected greater 
confidence in the pumped-slurry process and a saving in handling cost at 
the borrow pit. 

The stockpiled sand was bulldozed into a hopper and fed onto a conveyor 
where it was weighed as it was being delivered to a mixing tank constructed of 
concrete. Water from the deep-well pumps entered the tank through a series of 
jets that created the agitation necessary to maintain the solids in suspension. 
During most of the time water for slurrying was used at the rate of 5,000 to 
6,000 gpm. The sand, in slurry form, was taken from the tank through a suc- 
tion line at the rates varying from 200 to 350 tons per hour by a horizontal 
centrifugal pump, capacity 8,000 gpm. The slurry, 10 to 20 percent solids by 
volume, was impelled by a 650-hp diesel engine through 14-inch-ID pipelines to 
the various injection boreholes. The engine supplied sufficient energy to 
overcome the friction in up to 4,000 feet of pipe and deliver the slurry to 
the top of the borehole at a pressure to 50 psi when required. Figures 19 and 
20 are two views of the pumped-slurry mixing plant. 

Pressures at the head of each injection borehole were usually in the 
vacuum range (less than atmospheric), a characteristic of the pumped-slurry 
process, because of the drop to the mine level. These pressures would some- 
times rise momentarily above atmospheric pressure when the slurry pump was 
started at the beginning of a shift or when an occasional blockage occurred. 
Care was taken to stop the operation immediately when pressure increased dur- 
ing injection in a relatively shallow hole. 



38 




FIGURE 18. - View of earth-moving and screening equipment operations in the borrow pit. 




FIGURE 19. - View of end of hopper, conveyor, slurry pump, diesel engine, pipelines and con- 
trol building at mixing plant. 



39 




FIGURE 20. - View of mixing tank, suction line, sand being delivered to mixing tank, and 
bulldozer in the background pushing sand into hopper. 

Of necessity, the slurry distribution pipelines were buried beneath the 
busier downtown streets; this included all intersections. On the less- 
traveled side streets the pipelines were laid along the curb or the side of 
the road where arrangements were made with property owners to permit the con- 
struction of ramps over the pipelines at entrances to driveways. To facili- 
tate conveying slurry to injection boreholes in the downtown business district, 
it became necessary to lay the slurry pipeline under a spur of the Union 
Pacific Railroad at one point and under the mainline tracks at another. With 
permission of the railroad company, a 90-foot, 18-inch-diameter bore was pro- 
vided for this purpose under the spur track. The crossing under the mainline 
tracks was accomplished by installing the slurry pipeline through the existing 
pedestrian underpass. (See fig. 21.) 

In fiscal year 1975 and during Phase II, Congress appropriated an addi- 
tional $500,000 for the Bureau of Mines to conduct further backfilling opera- 
tions in Rock Springs, Wyo. This increase in funds provided an excellent 
opportunity for the Bureau to continue the backfilling of voids under and 
adjacent to critical area 11. It became apparent during Phase II that the 
voids under the southern and middle portions of critical area 11 were accept- 
ing far more sand than the calculations indicated. The excess was undoubtedly 
moving into old workings west of area 11 where surface support was also needed. 



40 




J&SL 



FIGURE 21. - View of the installed slurry pipeline in the pedestrian underpass beneath the 
mainline tracks of the Union Pacific Railroad. 

The additional funds, however, were not sufficient to' warrant contracting 
for a new backfilling project because at least $300,000 would be required to 
purchase and make operable the equipment comprising a new injection plant. It 
became expedient, therefore, to renegotiate the existing contract with WHAN 
Engineering and Construction, Inc. , and utilize the newly available funds 
almost exclusively for backfilling mine voids. 

On November 7, 1974, the contract was modified in the amount of 
$470,000.00, making the following negotiated changes and additions to the 
contract: 

1. The contractor shall continue to pump slurry into the coal mine voids 
at the contract unit price of $0.90 per ton until the quantity of material 
placed amounts to 175,000 tons (original contract). 

2. After the contractor accomplishes the work described in item 1 above, 
he shall then pump an additional estimated amount of 156,666 tons of slurry 
into the coal mine voids at a new unit price of $3.00 per ton. 



system. 



At the conclusion of the contract, as modified, the entire pipeline 
including all piping, both surface and underground, two bypass valves, 



41 



two squeeze valves, and the connector into the tub, which would have reverted 
to WHAN Engineering and Construction, Inc. , under the original contract shall 
become the property of the Bureau of Mines and all rights, title and interest 
in and to said pipeline shall vest in the Bureau of Mines free of all encum- 
brances and liens. 

Acquisition of the pipeline system compensated the Bureau to some extent 
for the increased unit cost per ton of injected material. Moreover, it was 
anticipated that the newly acquired pipeline system would be used by the 
Bureau in implementing a future backfilling contract at Rock Springs. 

This second contract modification increased the total contract amount from 
$680,167.00 to a new total of $1,150,167.00; increased the estimated injected 
fill material quantity from 165,000 tons to 331,666 tons; and increased the 
estimated project area from 28 acres to 54 acres to include most of area 11 
(see fig. 16). On May 29, 1975, the contract was modified a third time in the 
amount of $50,000 to make a new contract total amount of $1,200,167.00; whereby 
the estimated quantity of fill material injected was increased by an addi- 
tional 16,667 tons to make a new estimated total of 348,333 tons. 

When the project was completed June 26, 1975, an actual total of 
348,427 tons of solids had been injected into two project areas of the No. 1 
Seam mined coalbed. The total surface area overlying the backfilled portion 
of the coalbed was estimated to be approximately 55.2 acres. Six injection 
boreholes were used in the backfilling. Injection boreholes W-3 , W-5, and 
W-6 were located in inundated mine areas and boreholes W-l, W-2, and W-4 in 
dry areas. A total of 110,111 tons of sand was injected into dry voids. 

During the backfilling operation coincidental pothole-type subsidences 
occurred under and near the mainline tracks of the Union Pacific Railroad. 
These mine areas were not included in the original project because they were 
thought to have been isolated by rock-wall barricades constructed underground 
and effectively filled earlier by the mining company. Bureau engineers designed 
and coordinated a backfilling subproject between the Bureau contractor and the 
railroad company, whereby a 6-inch spur pipeline was constructed off the main 
14- inch-diameter slurry line to fill the isolated mine voids. Approximately 
5,300 tons of sand was successfully injected through a 1,200-foot plastic 
pipeline into nine 6- inch-diameter injection boreholes, filling the isolated 
mine voids to refusal with the water draining away through the old barricades. 
The progress of backfilling is summarized in table 3. 



42 



TABLE 3. - Injection data on the second large-scale demonstration 

project, 1974-75 



Project area and 


Injection 




Total 


acreage 


borehole 


Approximate injection periods 


injections, 
tons 


14, 8.20 acres 


W-l 


Aug. 7, 15-18, 21, and Sept. 17, 
1974. 


13,175 






Aug. 7-15, 18-19, 21, and 
Sept. 18-20, 1974. 


24,802 


11, 46.00 acres 


W-3 


Aug. 22-Sept. 14 and 
Sept. 21-22, 1974. 


62,023 






Oct. 3-9, 13-26, and Nov. 15-16, 
18-21, 1974. 


66,834 




W-5 


Sept. 28-0ct. 2, Oct. 26-Nov. 3, 
and Nov. 7-13, 1974. 


97,866 




W-6 


June 3-26, 1975 


78,427 


Union Pacific Rail- 


9 holes. . 


Oct. 13-25 and Nov. 8-9, 12, and 


5,300 


road, 1 acre. 




15-16, 1974. 





Backfill material was distributed in the mine workings throughout an area 
of approximately 55.2 acres, according to observations from the monitoring 
boreholes. As in the previous project the presence of fill material in moni- 
tor holes at levels higher than the roof indicated that filling of the mine 
workings within the 55.2-acre area was essentially complete from floor to roof. 
The only place within the confines of the planned project area that the exist- 
ence of fill is questionable was in dry mine voids along the north side of the 
railroad tracks between injection boreholes W-3 and W-4, where some suspected 
dead-ended updip mine rooms exist and/or mine workings are located updip from 
the injection boreholes in such a position that the gravity flow of the mate- 
rial prevented filling in the area. In addition the previously mentioned 
underground barricades referred to during the backfilling under the Union 
Pacific Railroad tracks may also have prevented filling in this small area 
(about 2 acres). Except for this, the estimated extent of the backfilled area 
is consistent with the estimated volume of void space in the mined bed and 
with the quantity of fill material that was injected. 

The cost of the hydraulic backfill demonstration project contract, in 
which 348,427 tons of sandfill material was injected, was $1,200,164.07. In 
addition, $34,842.70 ($0. 10/ton) was paid to Upland Industries Corp. for the 
acquisition of the sandfill material. The total cost of the project was 
$1,235,006.77 or $3.54 per ton. This can be compared with the $5.07 per ton 
cost of the previous large-scale demonstration project. 

i 

The demobilization of the operation, including cleanup and restoration of 
all project work areas, was conducted in accordance with the contract specifi- 
cations. Although it was situated on private land, restoration of the borrow 
pit was completed in accordance with the stipulations set forth by the Bureau 
of Land Management and included seeding of the borrow area in fall 1975 and a 



43 



verification inspection of the satisfactory growth after one complete growing 
season in fall 1976. 

PROJECT 3, WITH A MANAGEMENT SUPPORT CONTRACT 

In fiscal year 1976 Congress appropriated $1,500,000 for the Bureau of 
Mines to continue utilizing the pumped-slurry backfilling process in areas of 
Rock Springs, Wyo. , where surface support would be beneficial to the city. 

The sites selected for this, the third large-scale subsidence control 
project in the city, included critical areas 2, 9, 10, 12, 13, and 15 and the 
remaining portion of area 11. The objective of the project was to fill 
flooded and dry abandoned coal mine voids in the No. 1 and No. 7 Seams with 
solid material to support mine overburden and alleviate potential damage to 
surface structures, streets, and utilities. To accomplish this objective, it 
was estimated that approximately 350,000 tons of sand would be needed to fill 
the flooded and/or otherwise inaccessible mine voids beneath the critical 
areas enumerated above, a total area of about 90 acres. Figure 22 is a map of 
a portion of Rock Springs showing the location of the critical areas included 
in the project. 

In planning for the project, the Bureau decided that extensive drilling 
operations to provide monitoring could be dispensed with because the nature of 
the pumped-slurry filling process had been made apparent in earlier projects 
and was predictable in similar circumstances. Information as to the position 
and attitude of the coal seams was determined from the drilling of numerous 
boreholes. Depths to the mine voids ranged from 64 to 293 feet, with depths 
of alluvium ranging from 15 to 60 feet. Only seven boreholes encountered mine 
voids, and these were used for both injection and monitoring. The money saved 
in this manner was diverted to filling more of the mine voids. Appreciable 
savings would also accrue to the project because two water wells, two injec- 
tion boreholes, and some of the pipelines, which had reverted to the Bureau of 
Mines from the previous project (WHAN) , would be made available to the con- 
tractor for the third large-scale project. Moreover, because some of these 
facilities remained on property of the Union Pacific Railroad and in strategic 
position with respect to the project areas, the Bureau indicated in the con- 
tract specifications that the slurry injection plant, sand stockpile, and some 
of the slurry distribution pipelines be established in the same general area 
as in the previous project. In due course, the project contractor made satis- 
factory arrangements with the railroad company in order to comply with the 
specifications . 

Procurement of a contractor to do the specified work was again carried 
out under Federal Government competitive contracting procedures. On May 28, 
1976, a contract to do the work was awarded to M. J. Bober Co. , Littleton, 
Colo., in the amount of $1,043,650.00. This was increased by $1,353.00 to pay 
for the lease of the plant site on railroad property. Work on the contract 
was started June 4, 1976. Mobilization included preparation of a borrow pit, 
drilling to establish injection boreholes, and repair and modification of an 
existing water supply system, and pipelines. A mixing plant, slurry pumps, 
and recording equipment were also furnished and installed by the contractor. 



44 



&\ \K «&*2v /■■'-■'A 
ir><?S$8!fe fa$> 




FIGURE 22. - Map of a large portion of Rock Springs showing loca- 
tion of critical area 2 in the No. 7 Seam and critical 
areas 9, 10, 11, 12, 13, and 15 in the No. 1 Seam in 
relation to underground mine workings. 



45 



The sandfill material was obtained from a Bureau of Mines designated site 
located about 4 miles from the mixing plant on federally owned land controlled 
by the Bureau of Land Management (BLM) . This site was made available to the 
contractor through a "Free Use Permit" between the BLM and the Bureau, and had 
not been used as a source of material in any previous mine backfilling project. 
At the borrow site, the contractor furnished and installed screening and mate- 
rials handling equipment adequate to remove rock and debris in excess of minus 
5/8-inch size and produce sandfill at a rate of 2,800 tons per day. Shortly 
thereafter, under the terms of the contract, it was agreed to begin using pro- 
gressively larger screen sizes (up to minus 2-inch) to assess the capability 
of the process in handling sand with larger size rock particles and other 
debris. It was transported to the stockpile adjacent to the mixing plant site 
in 30- ton bottom-dump trucks. Removal of the fill material and subsequent 
reclamation of the borrow pit site were conducted in accordance with stipula- 
tions required by the BLM. 

The contractor purchased the installed deep-well pumps from the previous 
contractor (WHAN) and, after repairs and improvements, made operational a 
water-supply system using the existing wells and supply pipeline sufficient to 
deliver a minimum of 5,000 gpm to the designated mixing plant. The existing 
water well sites were located approximately 1,600 feet from the mixing plant 
site. The vertical distance from the top of the mine water pool in the No. 1 
coal seam to the mixing plant site was approximately 75 feet. The two wells 
were 135 feet deep into a 7-foot mine void and cased with 18-lnch-ID steel 
pipe to 95- and 98-foot depths. 

The mixing tank, which was basically the same as those constructed for 
the previous projects, was fabricated of steel and arranged in such a manner 
that it would receive measured and weighed sand at a rate of at least 
2,800 tons per day and water at a rate of at least 5,000 gpm. The water and 
sand were mixed to form a uniform slurry in the tank. Provisions were made 
to install the 12 -inch-diameter intake pipe of the slurry pump in such a sub- 
merged position that it would withdraw the mixed material (slurry) from the 
tank. The tank was adequate in size to not only handle the designed quanti- 
ties but also make sufficient allowance for surges and operational problems, 
thus helping to prevent overflows. 

A slurry pump was installed to withdraw the slurried sand and water from 
the mixing tank and impel it through the existing and constructed slurry dis- 
tribution pipelines to one of the two existing or one of the five new injec- 
tion holes. The pump was capable of moving a minimum of 2,800 tons of sand 
slurried with the required amount of water per day and was equipped in such 
a manner that the delivery rate of the pump could be regulated by a system of 
valves. The pump was designed to pump the required amount of slurry through 
a maximum distance of approximately 6,000 feet of pipeline to the various 
injection boreholes at a minimum velocity of 10 feet per second. It developed 
25 psi pressure at the top of each injection borehole when required. 

All operations involved in the mixing and injection processes were con- 
trolled from an instrument panel located in a building supplied by the con- 
tractor adjacent to the mixing tank. Metering devices recorded the following 



46 




FIGURE 23. - Map of a portion of Rock Springs showing the location of Project 3 
installations. 



47 



daily activity: total slurry pumping time, time to purge pipelines, gallons 
per minute of water and slurry pumped, tons of solid material injected, and 
pipeline pressures at the slurry pump and at the injection boreholes. Fig- 
ure 23 is a map showing the location of the project facilities. Figures 24, 
25, and 26 are views of the equipment that was being used at the borrow pit 
and at the slurry mixing plant. 

The fiscal funding constraints imposed by the Congressional appropriation 
for the Rock Springs backfilling work prevented the Bureau from providing its 
own personnel for managing and monitoring the contract activities beyond 
September 30, 1976. Because of this, the Bureau awarded a $15,000 contract 
to Johnson-Fermelia and Crank, Inc. , Rock Springs, Wyo. , on July 26, 1976, to 
provide the necessary daily onsite project management and monitoring services 
as required for the backfilling contract work in order to insure that the proj- 
ect work was being conducted in accordance with the contract specifications. 

The consulting firm of Johnson-Fermelia, and Crank, Inc. , possessed a 
working knowledge of the Rock Springs backfilling project work. The firm was 
located in Rock Springs in about the center of the project work area and had 
an excellent rapport with the city of Rock Springs, utility companies, Wyoming 
State Highway Department, Union Pacific Railroad, and all other necessary 
Federal, State, and local organizations. In addition, Wayne Johnson, a prin- 
cipal of the firm, was the City Engineer for Rock Springs and had been 
actively associated with the backfilling work since its beginning in 1970. 
As City Engineer, Johnson had approved all previous and current project work 
plans and had maintained for the city a close watch on all previous project 
work. 

On September 29, 1976, a second project management contract for $16,000 
was awarded to Johnson-Fermelia, and Crank, Inc. , to provide for necessary 
extended onsite project management and monitoring services to project 
completion. 

Specific monitoring services that were performed included, but were not 
limited to, the following: ' 

1. Intermittent or continuous daily inspection of the project work as 
directed by the Bureau representative and daily verification of completed work 
on the prime contractor's prepared daily work report sheets. 

2. Identifying and notifying the Bureau, including the authority to stop 
the work, of any potential project-related public health and safety hazard 
and/or any work being performed in violation of Federal, State, or local law 
or regulation. 

3. Performance of periodic borehole monitoring measurements with Bureau- 
provided equipment. 

4. Continuous inspection during the injection pumping operations, includ- 
ing performing periodic calibration checks of the fill material weighing scale. 



48 




FIGURE 24. - View of earth-moving and screening equipment being used in the borrow pit. 




FIGURE 25. - Photograph of sand being bulldozed into hopper and fed from conveyor into mix- 
ing tank. (Note jet action at side of tank creating turbulence.) 



49 









i 


*tt" 


5 




I j 


; p"-^uj 




tfBSyr 


g.... ■ 






/ir : : vfl^^E^ 


SB 


" 3H» 




" ? - ^^ 


*- ^B IBM 

i 


Ik- 




r 




^--j-^pv 


5£ 







K Z a 


.-' m '' ^■^fi'^*^^"^^ tct* 


:■■'>:.' ' 



FIGURE 26. - View of mixing plant operation: note suction line at bottom of mixing tank 
leading to slurry pump with pump motor in left foreground. 



5. Conveying Bureau work directives to the prime contractor as necessary 
and as directed by the Bureau representative. 

6. Taking photographs of the project equipment and operations represen- 
tative of the contract work progress. 

The monitoring services were performed by an engineering technician and/or an 
engineer with a vehicle as deemed necessary by the Bureau representative. 

The slurry operation began on July 15, 1976, and was completed on Octo- 
ber 12, 1976. A total of 397,464 tons of solids was injected into the seven 
separate areas of the No. 1 and No. 7 mined coalbeds. The total surface area 
overlying the backfilled portions of the coalbeds was estimated to be approxi- 
mately 90 acres, which included an expanded portion of area 11. Here, as in 
project No. 2, the sand was migrating into voids west of area 11, where it was 
known that surface support would also be needed. A total of seven injection 
boreholes, including two from the previous project, were used in the backfill- 
ing. Injection boreholes B-l, B-2, B-3, B-7, and B-8 in the No. 1 bed and B-6 
in the No. 7 bed were located in inundated mined areas, and borehole B-5 in 
the No. 1 bed was in a dry area. *A total of 378,712 tons of fill material was 
injected into inundated mine voids, and a total of 18,752 tons was injected 
into dry voids. The progress of backfilling is summarized in table 4. 



50 



TABLE 4. - Injection data on the third large-scale demonstration 

project, 1976 



Project area and 
acreage 



Injection 
borehole 

B-l 

B-2 

B-3 

B-6 

B-5.. 

B-7 

B-8 



Approximate injection 

periods, 1976 

July 15-25 

July 29 

Aug. 3-31 

Sept. 2-11, 13, and Oct. 8.. 
Oct. 4, 7-8 

July 30-31, Aug. 31-Sept. 2, 

and Sept. 13-0ct. 1. 
Oct. 5-6, 7-12 



Total injections, 
tons 



9, 17.70 acres. . 

10, 1.50 acres. . 
12, 3.30 acres. . 
2, 22.40 acres. . 
15, 10.50 acres. 

II. 1 20.10 acres. 



52,807 

3,176 

116,346 

41,262 

18,752 

114,194 



50,927 



x Expanded area 11. 



In order to provide pipeline access to areas 9 and 10, it was necessary 
to cross Wyoming State Highway 430 and Bitter Creek. Because of the antici- 
pated relatively brief time of injection into boreholes B-l and B-2 (15 days), 
the Wyoming Highway Department allowed the contractor to lay the slurry pipe- 
line across Highway 430 but required construction of a ramp over the pipeline 
so as to maintain the flow of traffic. In crossing Bitter Creek the city 




FIGURE 27. - View of slurry pipeline on the pedestrian foot bridge over Bitter Creek. 



51 



allowed the contractor to lay the pipeline on an existing pedestrian foot 
bridge, a less expensive and more convenient procedure than providing a sup- 
porting trestle (fig. 27). While the pipeline could have followed the ground 
in and out of the deep ravine, the number of vertical bends required would not 
only have increased the cost but also have added greatly to the friction head. 

On one occasion an accumulation of entrained air in the mine cavity 
exhausted with high velocity through injection borehole B-6 back through the 
open-ended pipeline at the pumping plant site. Although no damage resulted 
from the air discharge, it was concluded that in future projects a pressure 
release valve should be installed in the pipeline at the top of all the 
injection boreholes to avoid similar occurrences. 

When pumping through boreholes B-5 into rather shallow mine workings that 
were believed to be isolated from the main part of the mine, water was 
reported seeping to the surface near the pumping plant site. This water was 
determined to be the mine water that was being used for the backfilling opera- 
tion, and therefore pumping was stopped in borehole B-5 prior to the planned 
completion. 

In September, after the 350,000 tons of sand had been injected under the 
critical areas, it was noted that the mine voids, particularly in the north 
part of area 11, were accepting the sand slurry easily at less than atmos- 
pheric pressure. It was decided, therefore, to continue injection under the 
highly vulnerable downtown area until funds available to the Bureau were 
exhausted. This amounted to an additional $150,024.14, bringing the total 
cost of project work to $1,195,027.14. On October 12 the slurry pumping 
operation was completed, with a total of 397,464 tons of fill material having 
been injected into the old mine workings throughout an area of about 90 acres. 

A total of $29,959.00 was paid to the firm of Johnson-Fermelia and Crank, 
Inc., for providing the necessary onsite management and monitoring services, 
bringing the total cost of the project to $1,224,986.14, or $3.08 per ton. 
This compares with the $3.54 per ton cost of the previous or second large- 
scale demonstration project and $5.07 per ton for the first project. 

The demobilization of the operation, including cleanup and restoration of 
all project work areas, was conducted in accordance with the contract specifi- 
cations. Restoration of the borrow pit was completed in accordance with the 
stipulations set forth by the Bureau of Land Management and included seeding 
of the borrow area in fall 1976 and a verification inspection of the satis- 
factory growth after one complete growing season in spring 1978. 

CONCLUSIONS 

In positioning injection boreholes for the critical areas in the three 
large-scale projects, Bureau engineers assumed that the slurried backfill 
material would be distributed in the flooded mine workings more or less in 
equal distances from the points of injection. For dry mine voids, the 
injection boreholes were placed at higher levels in the workings to be filled. 
These assumptions were apparently realistic, according to observations from 



52 



32 monitor boreholes in the first project and from 20 monitor holes in the 
second project. Similar phenomena were noted during this period at the ini- 
tial project in Scranton (22) and in the laboratory model studies (appendix). 
The movement of fill material, therefore, having become predictable in the 
pumped-slurry process, monitoring for the third project was limited to those 
boreholes that were later used for injection. The estimated extent of the 
backfilled areas in the three projects, about 178 acres, was consistent with 
the estimated volume of void space/ in the mined beds and with the quantity of 
fill material that was injected. 

Of the methods of hydraulic backfilling formerly used, controlled flush- 
ing (see section of "Hydraulic Backfilling Methods") also results in well- 
filled mine openings because confinement is provided by bulkheads, fill 
placement is directed by hand into designated spaces, and the daily progress 
can be inspected. Controlled flushing and the pumped-slurry method are not 
generally competitive, however, because controlled flushing is limited to 
accessible mine workings. The alternative method of backfilling inaccessible 
mine openings, known as gravity blind flushing (see section on "Hydraulic 
Backfilling Methods") does not involve pumping of slurry and results in incom- 
plete filling, both laterally and vertically. Of the methods of hydraulic 
backfilling now known, therefore, the pumped-slurry technique (actually 
another form of blind flushing) , provides the most complete filling of mine 
workings that are flooded or otherwise inaccessible. 

The pumped-slurry method proved successful under the following conditions 
encountered in the three large-scale projects: 

Depth of mine workings to be filled, between 30 and 293 feet below the 
surface; 

Dip angle of workings, for the most part less than 6°; 

Mine workings relatively unobstructed by caving of overlying strata; 

Average depth of alluvium, 35 feet; 

Minimum rock cover, 5 feet; 

Particle size of screened sandfill material, minus 1/4-inch ranging 
up to minus 2 -inch; 

Specific gravity of sand particles, 2.6; 

Bulk density (dry), 100 pcf; and 

Water available in large quantities. 

Although the sand emplaced by the pumped-slurry process does not totally 
refill the space formerly occupied by the coal, it does support the remaining 
pillars and reduces the amount of breakage that otherwise would occur in the 
overlying strata. This lessens the chance that such disturbances might 



53 



eventually reach the surface and cause subsidence. There have been no 
reports of subsidence affecting surface areas overlying the backfilled areas. 
This included the Kerback-Belmont area where subsidence incidents had been 
continuing. 

Further use of the technique in different areas will define the range of 
conditions under which it is feasible. Modifications may extend the range of 
favorable conditions. The depth range for which the new method may be feasi- 
ble has not yet been defined. At shallow depth, material injected under pres- 
sure may rise to the surface rather than being confined to the mine level, 
especially in areas where overlying strata are fractured. The vertical com- 
pleteness of fill in mines that are well above water level needs to be deter- 
mined. The optimum size range of solids for efficient transport will be 
defined by future experimentation. 

The actual injection operations in the gravity blind flushing method 
require extensive drilling operations to provide injection boreholes and con- 
tinues truck traffic through city streets to bring fill material to injection 
boreholes. Similar disturbances, but to a lesser degree, are created when the 
controlled method is used. At the injection borehole, personnel are working 
in the street to direct the solids and water down the borehole. In the pumped- 
slurry method, street disturbance within the project area is limited to the 
drilling of an occasional borehole and the installation, maintenance, and 
removal of the distribution pipeline. During the second and third projects, 
the sand stockpiles, slurry mixing plants and pumps, as well as parts of the 
pipelines, were situated on the right of way of the Union Pacific Railroad. 
While strategically located with respect to the injection boreholes, the noise 
and dust associated with the operation of the plants was isolated from most 
dwellings. In the first project the slurry mixing plant and pumps were 
located at the borrow pit, approximately 2 miles from the built-up area, and 
had minimal impact upon the environment. During the 3-year period the three 
projects were implemented, the slurry moved quietly through the built-up areas 
in pipelines, many of which were buried. Aside from the drilling of boreholes 
and the installation and removal of pipelines, the only work required during 
the injection periods was cleaning the streets after infrequent pipeline leak- 
age and on a few occasions where slurry rose to the surface in uncapped or 
improperly sealed monitoring boreholes. 

Cost comparisons of the different methods of backfilling are difficult to 
make because the total number of projects span an inflationary period of 
rapidly rising costs. Moreover, subsurface conditions vary in the number of 
coalbeds to be filled, their depth, the thickness mined, and the percentage of 
coal left as pillars. Of four subsidence control projects in the Anthracite 
region backfilled entirely or mainly by the controlled flushing method between 
1963 and 1968, the cost per cubic yard or per ton of solids injected ranged 
from $1.84 to $2.38. On the average, a cubic yard of anthracite refuse weighs 
1 short ton. For two blind flushing projects in 1965 and 1967, the costs were 
$2.46 per ton — a cost that is extremely high when the limited effectiveness of 
the gravity blind flushing method is considered. In four projects between 
1966 and 1969 in which controlled and blind flushing methods were combined, 
the overall cost per ton ranged from $3.64 to $6.76. 



54 



The cost of the first large-scale test of the pumped-slurry method (at 
Scranton in 1972-73), in which about 451,000 cubic yards of crushed refuse 
was injected, was $2,165,915 — a unit cost of $4.80 per ton. At Rock Springs, 
where crushing cost was not included, costs per ton varied from $5.07 in the 
first project (only 152,467 tons), and $3.54 in the second (348,427 tons), to 
$3.08 per ton in the third project (397,464 tons). Absence of monitor bore- 
holes and reuse of some equipment undoubtedly account for part of the lower 
unit cost of the third project. 



55 



REFERENCES 

1. Ash, S. H. , and J. Westfield. Backfilling Problem in the Anthracite 

Region as It Relates to Conservation of Anthracite and Prevention of 
Subsidence. BuMines IC 7342, 1946, 18 pp. 

2. Ashmead, D. C. Mining of Thin Coal Beds in the Anthracite Region of 

Pennsylvania. BuMines Bull. 245, 1927, 113 pp. 

3. Candeub, Fleissig and Associates, Newark, N. J. Demonstration of a 

Technique for Limiting the Subsidence of Land Over Abandoned Mines. 
June 1971, 65 pp. (and appendix). Sponsored by Urban Renewal Agency, 
City of Rock Springs, Wyo. ; available from National Technical Informa- 
tion Service, Springfield, Va. , PB 212 708. 

4. Cochran, W. Mine Subsidence — Extent and Cost of Control in a Selected 

Area. BuMines IC 8507, 1971, 32 pp. 

5. Donner, D. L. , and R. H. Whaite. Investigation of Subsidence in Rock 

Springs, Sweetwater County, Wyoming. Unpublished report, September 
1969, 14 pp. Available for consultation from the Bureau of Mines, 
Division of Mining Environmental Technology, Washington, D.C. 

6. Enzian, C. Hydraulic Mine Filling — Its Use in the Pennsylvania Anthra- 

cite Fields, a Preliminary Report. BuMines Bull. 60, 1913, 77 pp. 

7. Griffith, W. , and E. T. Connor. Mining Conditions Under the City of 

Scranton, Pa. , Report and Maps. BuMines Bull. 25, 1912, 89 pp. (and 
separate case containing 25 maps) . 

8. Holmes, J. A. First Annual Report of Director of Bureau of Mines. 1911, 

pp. 42-43. 

9. Johnson--Fermelia & Crank, Inc. Underground Mining Study and Evaluation. 

Prepared for the Rock Springs Urban Renewal Board. Mar. 1, 1972, 27 pp. 
and appendices. 

10. Keith, R. E. Rock Springs and Blair Formations on and Adjacent to the 

Rock Springs Uplift, Sweetwater County, Wyoming (in Sedimentation of 
Late Cretaceous and Tertiary Outcrops, Rock Springs Uplift). Wyoming 
Geol. Assoc. Guidebook, 19th Field Conf. , 1965, pp. 42-53 (geologic map 
and isometric diagram of Rock Springs Fm. in pocket) . 

11. Kennedy, T. F. , and C. Hess. Report of General Investigation of Mining 

Conditions Under the City of Scranton, Lackawanna County, Pennsylvania. 
Report to Mayor and City Council of Scranton, June 14, 1960, 27 pp. 

12. Llghtfoot, W. E. Hydraulic Filling in Metal Mines. Calif. Dept. Conserv. . 

Div. Mines, Spec. Dept. 12, 1951, 28 pp. 



56 



13. Morgando, F. P. Engineering Geological Investigation of a Subsided Area, 

"D M Street — Connecticut Avenue, Rock Springs, Wyoming. Wyo. Highway 
Dept. Project ARS 1435, Mar. 28, 1969, 7 pp. (8 drillhole logs , 5 figs. ) . 

14. . Rock Springs Mine Backfill Project, Rock Springs, Wyoming. Wyo. 

Highway Dept. Project ARS 1523, Jan. 21, 1971, 2 pp. (9 drillhole logs, 
map) . 

15. Paone, J. , J. L. Morning, and L. Giorgetti. Land Utilization and Reclama- 

tion in the Mining Industry, 1930-71. BuMines IC 8642, 1974, 61 pp. 

16. Pennsylvania Anthracite Subsidence Commission. Final Report Submitted to 

the Legislature. Mar. 15, 1943, 12 pp. 

17. Pennsylvania Subsidence Committee. Report of the Subsidence Committee to 

the General Assembly of the Commonwealth of Pennsylvania. Mar. 1, 1957, 
54 pp. 



18. Rice, G. S. , and I. Hartmann. Coal Mining in Europe. BuMines Bull. 414, 

1939, 369 pp. 

19. Rock Springs, Wyo. Community Renewal Program. A Recommended Program for 

the City. July 1972, 32 pp. and appendices. 

20. Schultz, A. R. The Southern Part of the Rock Springs Coal Field, Sweet- 

water County, Wyo. Paper in Contributions to Economic Geology, Part II, 
Mineral Fuels. U.S. Geol. Survey Bull. 381, 1910, pp. 214-281. 

21. U.S. Bureau of Mines. Surface Subsidence Control in Mining Regions. 

Final Environmental Statement, FES 76-58, November 1976, 90 pp. 

22. Whaite, R. H. , and A. S. Allen. Pumped-Slurry Backfilling of Inaccessi- 

ble Mine Workings for. Subsidence Control. With an Appendix on Hydraulic 
Model Studies for Backfilling Mine Cavities by E. J. Carlson, Bureau of 
Reclamation, Denver, Colo. BuMines IC 8667, 1975, 83 pp. 

23. Works Progress Administration. Report on Preliminary Mine Flushing Sur- 

vey Under the City of Scranton, Pennsylvania. W. P. A. Work Project 
No. 1-35-80, Official Project No. 65-23-2614 (Proj. No. 4073), 
Feb. 27, 1936. 



APPENDIX 
REC-ERC-75-3 

HYDRAULIC MODEL STUDIES FOR 
BACKFILLING MINE CAVITIES 
(Second Series of Tests) 



by 

E.J. Carlson 



March 1975 

Prepared for 

U.S. BUREAU OF MINES 

under Modification No. 1 to Agreement No. H0230011, 

dated October 4, 1972 



Hydraulics Branch 
Division of General Research 
Engineering and Research Center 
Denver, Colorado 



UNITED STATES DEPARTMENT OF THE INTERIOR * BUREAU OF RECLAMATION 



ACKNOWLEDGMENT 

The studies described in this report were conducted by the author. 
Danny L. King is Chief of the Hydraulics Branch and J. C. Schuster is 
Head of the Hydraulics Research Section. Russell A. Dodge assisted in 
operating the model, taking data, and drafting the figures. The 
photographs were taken by Photographer W. M. Batts. Representatives 
from the Bureau of Mines, Denver, Colo., visited the Engineering and 
Research Center during the study to observe the hydraulic model tests 
and give guidance. A Memorandum of Understanding between the 
Bureau of Mines and the Bureau of Reclamation, Modification No. 1 to 
Agreement No. H0230011, was the contractural arrangement for the 
study. 



CONTENTS 

Page 

Purpose 1 

Summary and Conclusions 1 

The Model 1 

Model Box— Pump and Slurry Sump 1 

Piping and Measuring System 2 

Model Scales— Mine Pillars 2 

Backfill Material 2 

The Investigation 2 

Sloping Mine Floor 2 

Distorted model tests 2 

Undistorted model tests 3 

Blind entries— submerged mine 4 

Level Mine Floor 4 

Roof falls and cavities over roof falls— submerged mine 4 

Dry mine cavities 5 

Submerged mine cavities 5 

References 6 

Table 1 Summary Data of Model Tests 7 



LIST OF FIGURES 



Figure 



1 Model Test Facility 9 

2 Location of piezometers in mine roof and injection pipe 10 

3 Size analysis and relative density of backfill material 11 

4 Test 2— Contours of deposited backfill material 

(preliminary test— distorted model scale) 12 

5 Test 2— Photograph of deposited backfill material 

(preliminary test— distorted model scale) 12 

6 Test 3— Contours of deposited backfill material 

(preliminary test— distorted model scale) 13 

7 Test 4— Contours of deposited backfill material 

(preliminary test— distorted model scale) 14 

8 Test 4— Closeup photograph of deposited backfill material 

(preliminary test— distorted model scale) 14 

9 Test 5— Continuation of test 4— Contour map of deposited 

backfill material (preliminary test— distorted model 

scale) 15 

10 Test 5— Photograph of deposited backfill material 

(preliminary test— distorted model scale) 15 

11 Test 7— Contours of deposited backfill material 16 

12 Test 7— Photographs of pillar arrangement and deposited 

backfill matieral at end of test 17 



LIST OF FIGURES-Continued 
Figure Page 

13 Test 8— Contours of deposited backfill material at end 

of test showing the breakout channel 18 

14 Test 8— Photographs of deposited backfill material 19 

15 Test 9— Photographs of deposited backfill material — First 

in a series of three tests with blind entries 20 

16 Test 9— Location of blind entries and contours of 

deposited backfill material 21 

17 Test 1 1— Photograph of deposited backfill material— Last 

test in the series of three tests 21 

18 Test 1 1— Contours of deposited backfill material at the end 

of the series of three tests 22 

19 Test 1 1 —Deposit pattern above 3-foot contour with blind 

entries 23 

20 Test 17— Mine pillars, simulated roof falls, and cavities 

in mine roof before test started 24 

21 Test 17— Schematic section for tests with roof falls and 

cavities over roof falls 25 

22 Test 17— Photographs of backfill deposit at end of 

test— Roof falls and cavities over roof falls were simulated 26 

23 Test 17-Contours of deposited backfill material with 

roof falls and roof cavities in the mine 27 

24 Test 18— Contours of deposited backfill material — Roof 

falls and roof cavities over roof falls in the mine 28 

25 Test 12— Contours of deposited backfill material in the 

first of a series of four tests in dry cavities 29 

26 Test 12— Photographs of deposited backfill material at 

the end of the test in a dry cavity— First in a 

series of four tests 30 

27 Test 13— Second in a series of four tests in a dry 

cavity— Contours of deposited backfill material 31 

28 Test 13— Closeup and overall photographs of deposited 

backfill material at the end of the test 32 

29 Test 14— Contours of deposited backfill material- 

Third in a series of four tests 33 

30 Test 14— Photographs of deposit pattern around injection 

hole in a dry cavity— Third in a series of four 

tests 34 

31 Test 15— Contours of deposited backfill in a dry cavity 

at the end of the series of four tests 35 

32 Test 15— Photographs of deposited backfill material at 

the end of the series of four tests in dry mine cavities 36 

33 Test 16— Contours of deposited backfill material at the 

end of the test in a submerged cavity 37 

34 Test 16— Photographs of backfill deposit at 

the end of the test showing the final breakout channel 38 



PURPOSE 

This study is a continuation of the investigation 
reported in Report REC-ERC-73-19, "Hydraulic Model 
Studies for Backfilling Mine Cavities." The Bureau of 
Mines asked the Bureau of Reclamation to conduct 
additional hydraulic model tests to study different 
aspects of backfilling mine cavities with sand and waste 
material to reduce subsidence of land at the surface 
above the mine cavities. 



SUMMARY AND CONCLUSIONS 

Additional tests were made in the model of an 
idealized coal mine that was operated to determine the 
results of various conditions where mine cavities are 
backfilled by pumping a fine sand slurry. Fine, uniform 
blow sand having a median size of 0.14 millimeter 
obtained from the Rock Springs, Wyoming area was 
used to produce the sand slurry. 

Eighteen tests were conducted in this second phase of 
hydraulic model tests. The following mine conditions 
with slurry injection were simulated: 

1. Sloping floor with cavity submerged 

2. Level floor with cavity submerged 

3. Level floor with cavity dry 

4. Simulated mine with and without blind entries 

5. Corridors and rooms in which there were roof 
falls and cavities in the roof over the roof falls 

Conditions under which the tests were made are 
summarized in table 1. 

Conclusions from the first series of tests were reported 
in Report REC-ERC-73-19. Data from the second 
series consisting of 18 tests lead to the following 
conclusions which are in addition to conclusions made 
for the first series of tests. The results from the 18 tests 
reported here support the conclusions derived from the 
first series. 

1. As deposited backfill material reaches the 
quantity and pattern to build up back pressure in 
the injection system, one final breakout may occur. 
A channel is formed down an unobstructed corridor 
between rows of pillars. The entire discharge from 
the injection pipe goes down this one channel with 
high enough velocity to keep the fine sand moving 
without causing a high back pressure in the 



pipeline. This final breakout channel can transport 
slurry material for a long time over a comparatively 
long distance away from the injection pipe. Deposit 
would occur at each cross channel junction between 
the pillars. 

2. Fine sand backfill material injected into a sub- 
merged mine is transported over roof falls that 
block corridors when there is an open cavity over 
the top of the roof fall. In the model study, backfill 
material almost filled the cavities above the roof 
falls at the end of the tests. 

3. The extent to which backfill material will be 
transported into and deposited in slack water areas 
(blind entries) will depend on the position and 
geometry of the entry with respect to slurry flow 
past the entry. Fill material will deposit in slack 
water areas (blind entries) if circulation of sediment- 
laden water occurs in and out of the slack water 
areas. 

THE MODEL 

Model Box— Pump and Slurry Sump 

A watertight box made from 3/4-inch waterproof 
plywood, 15 feet square and 2-1/2 feet deep, was used 
to contain the model, figure 1. This box was used for 
previous tests of backfilling mine cavities as described 
in Report REC-ERC-73-19 prepared for the Bureau of 
Mines. The same slurry sump and 2-1/2-inch Kimball- 
Krogh sand pump as described in that report was used 
to pump the slurry from the sump to the model mine. 
A recirculating system was used in which fine sand was 
mixed with water in a sump that was 8 feet long by 2 
feet wide by 3.5 feet deep mounted below the floor. 
Slurry material was pumped into the center of the 
model, in every case, through the injection pipe 
mounted in the removable mine roof. Sand material 
was deposited in the mine and water would flow to the 
simulated water table. The water level was held above 
the mine cavity for submerged cavity tests. For dry 
cavity tests, the water level control gate on the model 
box was lowered completely; water would leave the 
elevated mine floor, flow into the model box surround- 
ing the section of simulated mine, out the 4-foot-long 
by 2.5-foot-wide sluice channel, and back to the slurry 
sump. Water level in the model box and slurry sump 
was maintained at the desired level by adding water as 
necessary. With the propeller mixer running at constant 
speed, concentration of fill material in the slurry sump 
would be varied according to the level of energy 
imparted to the fluid slurry by the mixer and according 
to the depth of fill material deposited in the slurry 
sump. 



Piping and Measuring System 

Previous tests showed that the general pattern of 
deposit was not dependent upon slurry concentration 
nor on injection pipe velocity, providing velocities were 
high enough to transport sediment without deposition 
in the injection pipe. The concentration and pipe 
velocities in the tests described here, therefore, were 
not intended to duplicate those conditions of the Rock 
Springs injection operations. The deposits should pre- 
dict the pattern that would occur in a typical mine 
with a symmetrical uniform pattern of mine pillars and 
cavities. 

For tests 1 through 3 the vertical intake to the sand 
pump used in previous tests was left in place. 
However, the pipe entrance was about 3 feet away 
from the vertical mixer propeller. To obtain a more 
uniform slurry concentration, the 2-inch nominal 
intake pipe was lengthened and set on a 45 angle to 
the vertical so the intake would be closer to the 
propeller mixer. The propeller mixer was used to keep 
the fine slurry sand in suspension. A 1/2 -inch feed pipe 
was used for tests 1 through 6 ancr replaced by a 
3/4-inch pipe for the remainder of the tests described 
in this report. The Venturi meters used for measuring 
discharge in previous tests were removed. A 3/4-inch 
Annubar flowmeter with 0.824-inch inside diameter 
was installed in the horizontal section of the pipe for 
measuring discharge in all tests, figure 1. To minimize 
possible plugging of the impact and low-pressure ports 
in the flowmeter, two purge water lines were attached 
to the ports, each line having a rotameter to measure 
the purge water. 

The Annubar flowmeter was calibrated for clear water 
without the purge inflow at the pressure ports. Flows 
through the rotameters were then set to give the same 
discharge rating as without the purge water connec- 
tions. To continuously determine the slurry discharge 
without getting fine sand in the meter ports and 
plugging them, a small amount of purge water was 
used. 

Pressure piezometers were located about one pipe 
diameter from the end of the injection pipe and in the 
mine cavity as shown on figure 2. The pressures were 
read on the water manometer board and recorded at 
short time intervals to determine changes in the 
pressure as the backfill material deposited in the cavity 
during each test. For tests 1 through 5, seven 
piezometers were used including a piezometer showing 
the water table elevation surrounding the mine cavities. 
For tests 6 through 17, pressures were measured at 11 
points. At points where fill material deposited up to 
the mine roof, the pressure taps became plugged with 
the fine sand. 



All figures in the report showing drawings and contour 
maps of the mine model are oriented with north at the 
top of the figure for easy comparison of the deposit 
patterns. Contour intervals are designated in feet above 
the mine floor on all contour maps. 

Model Scales— Mine Pillars 

The model mine was constructed to represent a mine 
with the cavity volume equal to 60 percent and the 
pillar volume equal to 40 percent of the total volume. 
The horizontal scale for all model tests was 1 m :48 p (1 
in the model is equal to 48 in the prototype.) The 
vertical scale for most tests were also 1 m :48 p . Some 
early tests (1 through 5) were made with a vertical 
scale of 1 m :14.908 p , a vertical distortion of 3.22, to 
establish deposit patterns with velocities in the model 
mine cavity equal to the velocities in the typical 
prototype cavity. Deposit patterns for undistorted and 
distorted scales were similar; therefore, tests 6 through 
18 were performed with the model constructed to an 
undistorted scale of 1 m :48 p , vertical and horizontal. 

Mine pillars were constructed in the model to represent 
horizontal dimensions 40 feet long and 10 feet wide, 
with a cavity spacing of 10 feet between sides of pillars 
and also 10 feet between ends of pillars. This gave a 
mine arrangement as described above with 40-percent 
solid and 60-percent cavity both for the distorted and 
undistorted model scales. The 8-foot-square mine area 
in the model represented a 384-foot square or 3.39 
acres in the prototype. 

Backfill Material 

Fine sand obtained from the Rock Springs injection 
project was used in the model studies. A size analysis 
and relative density determination for the backfill 
material used in the model and prototype mine is 
shown in figure 3. The median diameter of the fine 
sand was 0.14 millimeter. Standard properties and 
bearing capacity tests on the backfill material were 
made in the Soils Laboratory of the Earth Sciences 
Branch of the Bureau of Reclamation. These studies 
are reported in Report REC-ERC-73-19. 



THE INVESTIGATION 

Sloping Mine Floor 

Distorted model tests. -Tests 1 through 5 were made to 
evaluate the changed piping system, the Annubar 
flowmeter, the seal of the roof against the mine pillars, 
and general operation of the pump-piping system and 
slurry sump. Tests 1 to 3 were conducted with the 
vertical intake pipe on the pump. At the end of test 3, 



an inspection showed a hard crust of fine sand in the 
slurry sump just below the vertical pipe intake located 

3 feet horizontally from the mixer propeller. The crust 
which apparently formed over a period of operation, 
was similar to hard surface crusts that form in open 
channels having bed material made from fine sand. 

In test 1 sand was fed to the slurry sump at a rate of 
1.1 pounds per minute. For test 2 the rate of sand was 
increased to 12 pounds per minute. On both tests 1 
and 2, pressure built up in the mine cavity after fill 
material was deposited up to the roof level. In test 2, 
the pressure increased so much that the roof lifted 
from the pillars, and fill material was transported 
between the roof and the tops of the pillars, figures 4 
and 5. After test 2 was completed, four bolts were 
installed through the pillars from the mine floor to the 
roof. An additional 1/4-inch layer of sponge rubber 
was fastened with adhesive to the roof to form a seal 
on the pillars as the bolts were tightened. After 
conducting test 2 for about 35 minutes in the model, 
the mine cavities were filled to cause back pressure in 
the injection pipe. Shortly after this pressure built up, 
breakouts occurred upslope first, then downslope and 
to the slides. The mine was set on the 5 dip. 

Test 3 was operated for about half an hour. Water 
discharge was started at 0.030 ft 3 /s. When sand was 
added to the slurry, the discharge dropped to 0.020 
ft 3 /s. The average discharge during the test was 0.025 
ft 3 /s. Pressures measured at the piezometer on the end 
of the injection pipe varied from 1.02 to 1.74 feet 
compared to 1.56 to 1.74 feet measured at the 
piezometers in the cavity. Figure 6 shows the deposit 
pattern at the end of test 3. 

After completing test 3, the 2-inch intake pipe was 
lengthened 1 foot 3 inches and was reconnected to the 
pump intake at a 45 angle. With this arrangement, the 
end of the intake pipe was 0.9 foot above the floor of 
the slurry sump and closer to the mixer propeller. Test 

4 showed that moving the intake pipe closer to the 
mixer propeller caused extra deposit in a cone shape 
around the mixer in the slurry sump. Consequently, 
much of the sand added at 14 pounds per minute 
deposited in the slurry sump and was not pumped to 
the model mine. The amount of fill material that was 
pumped and deposited in the mine was comparatively 
small, figures 7 and 8. Test 5 was therefore made as a 
continuation of test 4. 

Tests 4 and 5 were made with the mine submerged and 
dipping 5°. The model had a horizontal scale of 
1 m :48 p with a vertical distortion of 3.22. The break- 
out through the initial deposit caused clouds of slurry 
to come through the corridors to the edge of the mine 



area. Pressures on the end of the discharge pipe in test 
4 remained approximately the same throughout the 
test, indicating that there was no back pressure; 
consequently, the solid fill material did not fill the 
cavity near the ceiling. In test 5, pressure at the end of 
the discharge pipe increased with continued injection. 
Figures 9 and 10 show the deposit pattern in the mine 
cavity at the end of the test. 

Undistorted model tests.— For tests 6 through 18 the 
pillars were changed to give an undistorted geometric 
scale of 1 m :48 p in both horizontal and vertical 
directions. An observation test was conducted with the 
mine submerged and dipping at an angle of 5 . 
Velocity in the 1/2-inch pipe was about 9.9 ft/s. 
Deposits occurred and silty water that could be 
observed at the edge of the mine section was moving 
upstream in corridors 3 through 9, counting from the 
left side looking downslope. Additional piezometers in 
the mine roof were added to give a wide pattern of 
pressure distribution away from the injection pipe. As 
the backfill material deposited, pressures with the 
additional deposit ring were slightly higher than pres- 
sures outside the central cavity (piezometers 7 through 
11). The additional piezometers 7 through 11 were 
added in the mine roof after test 6 was started. The 
test was stopped, the mine drained, the roof was raised, 
and the piezometers installed. No photographs were 
taken nor was a contour map of backfill deposit 
prepared for test 6 because of the changes during 
testing. 

Test 7 was performed with the same conditions as for 
test 6 except the injection pipe with an inside diameter 
of 1/2 inch was replaced by an injection pipe with an 
inside diameter of 3/4 inch to get higher discharge 
capacity through the pump-piping system. The 1/2- 
inch pipe was restrictive, which caused debris to collect 
in the pipeline. A valve was installed on the high point 
to the bowl of the centrifugal pump which made it 
possible to bleed air and later to extract sediment 
samples from the pump. The valve also made it easier 
to prime the pump at the startup for a test. At the end 
of test 7, material was flowing between the pillars and 
the mine roof in a few places, figures 1 1 and 12. After 
test 7 was completed, two additional toggle bolts, 
making a total of 6, were installed to hold the roof 
tight against the pillars. 

A water purge system for the Annubar flowmeter was 
installed at the end of test 7. Previous test discharges 
were set with only water in the piping system before 
the mixer was turned on. Without the purge system, 
when fill material was pumped in slurry form, the ports 
to the flowmeter would tend to plug. By using the 
purge system, pressure was positive at each of the two 



ports of the Annubar flowmeter, which caused a small 
flow into the pipeline, preventing fine sand from 
entering and plugging the pressure tubes to the 
flowmeter. 

Test 8 was made with a velocity of approximately 7.5 
ft/s in the 3/4-inch injection pipe. The mine was 
dipping 5° and submerged. The deposit pattern was 
observed at the end of the test after the roof was 
raised, figures 13 and 14. No deposit on the top of the 
pillars indicated the roof held tight against the back 
pressure that occurred in the mine cavity. After the 
initial deposit ring was established around the injection 
pipe, back pressure built up and a breakout occurred 
downslope in corridors 8 and 9, counting from the left 
looking downslope. The fine sand was carried in 
suspension along the bed and deposited in a large 
mound off the edge of the mine platform. With 
increased pressure, a breakout occurred and high 
velocity flow started upslope in corridor 9, counting 
from the left looking downstream. 

The characteristics of test 8 were typical of an 
injection into a submerged cavity with, open corridors. 
After the initial deposit ring has occurred and back 
pressure builds up in the cavity and in the injection 
pipe, a breakout occurs in one or two corridors. Fill 
material is carried along this channel in suspension or 
as bedload according to basic sediment transport 
principles. With the full flow of the injection pipe 
discharging along a channel, an equilibrium condition 
develops for sediment transport. Fill material deposits 
at intersections to essentially block side corridors and 
confine the flow along the one channel. Deposit builds 
in the channel until the cross-sectional area reduces and 
the velocity increases to cause critical transport 
conditions. 

Reports from field operations at Rock Springs, 
Wyoming, indicate flow occurs in a single channel over 
long distances after fill material is deposited up to roof 
level around the injection hole. Model tests showed 
that flow in a single breakout channel started when 
deposit sealed or nearly sealed the space adjacent to 
the roof around the injection hole. Pressure would 
build up in the cavity prior to the breakout and would 
lower as flow started in a single channel. Extensive 
deposit and lowered pressure prevented other breakout 
channels from forming. For test 8, after material had 
deposited up to the roof, slurry flowed down one 
corridor until the test was stopped. 

Blind entries— submerged mine.— Tests 9, 10, and 11 
were made as a series with the mine dipping 5° and 
submerged. After each test was stopped, the roof was 
raised and the deposit pattern observed. The mine roof 



was then lowered, fastened in place, and the next test 
in the series continued. A contour map and photo- 
graphs were made of the deposit pattern for tests 9 and 
11. Blind entries were simulated in the model by in- 
stalling blocks at various places in corridors in the 
mine. Some blocks were installed to block corridors at 
ends of pillars and also near the middle area of pillars. 
Some blocks were installed to prevent communication 
within corridors and over considerable distances in 
some cases, figures 15 through 19. During test 9, initial 
fill material deposited up to the roof around the injec- 
tion pipe, and back pressure caused a reduction in dis- 
charge. Piezometers attached to the roof showed the 
increase in back pressure and then the sudden decrease 
in back pressure when a breakout occurred. The 
pictures and contour map prepared at the end of test 
9 are shown in figures 15 and 16. 

Test 10 was a continuation of test 9, using a smaller 
discharge. The smaller discharge resulted in a lower 
intake velocity and, consequently, a lower sand con- 
centration. A small additional amount of fill material 
was deposited in the mine during test 10. The pressure 
in the area around the injection hole was comparatively 
high. No photographs were taken nor was a contour 
map prepared at the end of test 10. Before test 1 1 was 
started, the sand deposit was carved back to the 
deposit pattern left at the end of test 9. 

Test 11 was made with a slightly lower average 
discharge throughout the test. When the discharge 
tended to decrease because of back pressure, the 
control valve was opened to maintain a constant 
discharge. For test series 9, 10, and 11, the fill material 
seemed to deposit downslope first, then upslope, and 
then on the level out from the injection hole toward 
the sides. At the end of test 11, the last breakout 
established a comparatively high velocity flow upslope 
in corridor 8, counting from the left side looking 
downslope. The flow being confined to a single 
corridor caused the velocity to be comparatively high 
and, thus, the transport capacity continued at a 
comparatively high value. The flow at the end of test 8 
for a mine without blind entries was similarly confined 
to a single corridor. At the end of the series of tests 9, 
10, and 11, flow was confined to a single corridor and 
the slurry traveled upslope with a comparatively high 
velocity. Figures 17, 18, and 19 show a photograph 
and contour maps of deposited slurry material at the 
end of test 1 1 . Figure 19 indicates that fill material will 
not enter and deposit where blind entries prevent flow 
circulation. 

Level Mine Floor 

Roof falls and cavities over roof falls— submerged 



mine.— Tests 17 and 18 were conducted to show how 
fill material would be transported over simulated roof 
falls and through cavities above the roof falls, a 
condition occurring in coal mines after being 
abandoned for some time. Roof falls were simulated by 
truncated wood pyramids sloped 60° from the floor, 
figures 20 and 21. The top of the roof falls were 6 feet 
above the mine floor, the same height as the normal 
roof. Above the roof fall a cavity was formed by 
cutting the roof and constructing a box over the hole 
cut in the roof, figure 20. A piezometer was placed in 
each simulated cavity to measure pressures developed 
in the cavities during the backfilling operation. Two 
roof falls were placed at intersections of corridors; one 
roof fall was placed between ends of pillars and one 
was placed between sides of pillars, figure 21. 

For tests 17 and 18, the mine was level and submerged. 
Before beginning test 17, sand was added to the slurry 
tank so that the backfill supply was sufficient to 
complete the test. During the 50-minute test, samples 
of slurry taken from the pump discharge pipe varied in 
concentration from 1.2 to 5.1 percent, by weight, with 
an average of 1.8. These samples were taken using a 
1/8-inch tube with its entrance pointing upstream in 
the vertical pipe where flow lines were parallel. A 
photograph showing backfill deposit at the end of test 
17 and contour maps at the end of tests 17 and 18 are 
shown on figures 22, 23, and 24. Test 18 duplicated 
test 17, except test 18 had a higher injection velocity 
and higher slurry concentration, table 1. At the end of 
both tests 17 and 18, high velocity flow moved along 
one corridor directly away from the injection pipe. The 
velocity along the breakout corridors was high enough 
to transport fill material without depositing. 

At the end of test 17, slurry was flowing up over the 
roof fall, through the cavity above, and down the 
corridor, figure 23. The mine was submerged and the 
resistance offered by the roof fall was not great enough 
to cause slurry flow to move to another corridor. The 
slurry takes the flow path of least resistance. 

At the end of test 18, the last breakout channel was 
along a corridor adjacent to a corridor having a roof 
fall at a corridor intersection, figure 24. This breakout 
channel was in the opposite direction from the last 
breakout channel for test 17. It is apparent that for a 
level mine that is submerged, there is very little 
difference in the resistance to flow in one direction 
than to the flow in the opposite direction. In tests 17 
and 18, the last breakout channels were along the 
length of the pillars. Apparently, the abrupt expansion 
and contraction losses caused by the intersections of 
lateral channels in the short direction of the pillars may 
be greater than the losses in the corridors along the 



length of the pillars. For dry mine cavities, the flow 
conditions and, consequently, the patterns of resist- 
ance to flow are different from those in submerged 
cavities. 

Dry mine cavities.- Tests 12 through 15 were con- 
ducted with the mine cavity in an unsubmerged (dry) 
condition. The water table is lower than the floor of 
the mine cavity. To provide for this condition in the 
model, the injection water was allowed to drain out of 
the model box by having the water level control gate 
completely lowered. The mine roof was in place for all 
four tests. Fill material is deposited and slurry water 
returns to the water table. The backfill material in a dry 
mine cavity develops a deposit with a surface slope that 
is dependent on the critical tractive force (To = yDS) 
for the material, where T D = tractive force, y= specific 

weight of water, D = depth of water flowing over the 
deposit, and S - slope of the flowing water surface. A 
critical tractive force (tractive force that causes a given 
size of fill material on the bed to start moving) is 
related to the depth and the slope so that the product 
(DXS) is constant for the given size bed material. 

Tests 12 through 15 were conducted as a series in 
which fill material was not removed at the end of each 
test, table 1. A photograph and contour map were 
made at the end of each test to compare the progress 
of fill deposit, figures 25 through 32. The progress of 
the deposit with each successive test can best be 
observed on the contour maps, figures 25, 27, 29, and 
31. The 1-, 3-, and 5-foot contours show the deposit 
buildup and how the sloping face of deposited backfill 
material moves with time. Backfill material builds up 
close to the roof near the injection pipe. A breakout 
occurs when deposit around the injection pipe is high 
enough to force most of the flow in one concentrated 
channel, figures 27, 29, and 31. The direction of the 
breakout channel varies for different tests when the 
mine floor is level, indicating initial deposits are 
uniform and symmetrical for a symmetrical pillar 
pattern on a level floor. The difference in resistance to 
breaking out in one direction compared to another 
direction is very small. 

Discharge and, consequently, injection pipe velocity 
for the series of four tests conducted in a dry cavity 
were very nearly identical, 0.013 or 0.012 ft 3 /s and 3.5 
or 3.2 ft 3 /s, respectively. Solids concentration in the 
injection pipe varied from 1.0 (test 15) to 4.5 percent 
(test 12) by weight, table 1 . 

Submerged mine cavities.- Test 16 was conducted in 
two parts over a period of 2 days on a level submerged 
mine cavity. The first part had a discharge of 0.013 



ft 3 /s with a slurry concentration of 0.74 percent by 
weight, and the second part had a discharge of 0.021 
ft 3 /s and a slurry concentration of 6.7 percent. 
Photographs, figure 34, and a contour map, figure 33, 
were made at the end of the test on the second day. 
The deposit pattern, particularly the mine area that 
had deposited material up to and very near the roof, 
was extensive. This was caused by the high concentra- 
tion and higher discharge and, consequently, higher 
injection velocity. The higher velocity caused the high 
concentration of backfill material to deposit at a 
greater depth farther from the injection pipe than 
could be obtained with a lower injection velocity. 

The breakout channel develops whether or not there is 
a high injection velocity or lower injection velocity. 
For a higher injection velocity, pressure buildup in the 
mine cavity caused by pumping and deposit of backfill 
material takes longer than for a lower injection 
velocity. The deposit depth up to the roof or near the 
roof extends over a greater area for higher injection 
velocities. 



Chicago, Illinois prepared for U.S. Bureau of Mines, 
January 1972 

Yalin, M. Selim, "Theory of Hydraulic Models," The 
MacMillan Press, Ltd., 1971 

Yalin, M. Selim, "Mechanics of Sediment Transport," 
Pergamon Press, 1972 

Graf, Walter H., "Hydraulics of Sediment Transport," 
McGraw-Hill Book Company, 1971 

Ackenheil, Alfred C, and Dougherty, Murray P., 
"Recent Developments in Grouting for Deep 
Mines," Journal of Soil Mechanics and Foundations 
Division, Proceedings of ASCE, Vol. 96, No. SM1, 
January 1970. 



In test 16, some cavities were left in corridors between 
pillars. Slurry material was transported past opposite 
ends of pillars depositing material at the same time 
from opposite ends of corridors. The deposit blocked 
the corridors, leaving a small unfilled cavity. In a 
prototype mine, the extent of unfilled cavities between 
pillars depends on the pattern of the pillars and 
corridors. Cavities could be left between pillars when 
the flow pattern was symmetrical. 

A final breakout channel on test 16 formed at the end 
of the test in which most of the injected slurry was 
flowing down one corridor. This type of flow would 
continue if the test were not stopped. The slope of the 
bottom of the deposit in this last breakout channel was 
very flat, similar to the final breakout channels in 
previous tests in a submerged mine. 



REFERENCES 

Carlson, E. J., "Hydraulic Model Studies for Backfilling 
Mine Cavities," Bureau of Reclamation, Report 
REC-ERC-73-19, prepared for U.S. Bureau of 
Mines, October 1973 



Whaite, Ralph H., Allen, Alice S., "Pump-slurry Back- 
filling of Inaccessible Mine Workings for Subsidence 
Control," U.S. Bureau of Mines, Washington, D.C., 
1975. 

Courtney, W. J., Singh, N. M., "Feasibility of Pneu- 
matic Stowing for Ground Control in Coal Mines," 
Illinois Institute of Technology Research Institute, 



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15 x 15 x 2.5 Deep box 



Flowmeter 
Manometer board 

Twin rotameter 
purge system 



8 x 2 x 3.5 Deep sediment 
sump box below floor 



Figure 1 . Model test facility. 







































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■ ! 



10 



"I n n 



•PIEZOMETER IN INJECTION PIPE 

LJ LJ l_l LJ L_l LJ U 
2 



BEFORE TEST 5 

I I I I I i 
AFTER TEST 5 




Hiok 
A. 



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10 X 40 X 6 PILLAR BLOCKS 




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PIEZOMETER 8 THROUGH II 
NSTALLED AFTER TEST 5 



Figure 2. Location of piezometers installed in mine roof and injection pipe. 



10 



7-1742 (8-70) 

Bureau of Raclomorion 



PHYSICAL PROPERTIES SUMMARY PLOT (Relative Density) 



HYDROMETER ANALYSIS 
TIME READINGS 



SIEVE ANALYSIS 



US STANDARO SERIES CLEAR SQUARE OPENING 




001 002 



010 037 074 149 297 42 590 119 2 2 36 476 9 52 191 311 

DIAMETER OF PARTICLE IN MILLIMETERS 



76.2 127 IS2 



f'"E 



SAND 



MEDIUM 



f"*l 



GRAVEL 



COAIISE 



130 


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- 



MMIMUM 
DIHSITT 
3CALI 



RELATIVE DENSITY •% 

SCALED TO PLOT AS STRAIGHT LINE 



Classification Symbol SM 

Gradation Summary 
Gravel Q. 

Sand 84- 

Fines / In 

Atterberg Limits 

Liquid Limit 

Plasticity Index NP 

Shrinkage Limit ___^ 

Specific Gravity 
Minus No. 4 2.(o8) 

Plus No. 4 

Bulk 

Apparent 

Absorption 

Relative Density 
Minimum Density 

( 

Maximum Density 

fj < 

In-place Density 



% 



8 5,0 pcf 

L 3to gm/cm3) 

/jQ3j±pcf 

114- gm/cm3) 

93.3 PCF 

( /SO gm/rm3) 

Percent Relative Density 43S 
Permeability Settlement 

Placement Condition 

Coef of Permeability ft/yr 

( cm/sec) 

Settlement Under 

psi Load % 

( kg/cm2) 



'o Notes:. 



100 

MAXIMUM 
DENSITY 
SCALE 



Ll as determined in laborator y 
y*tode/ shirty 



Sample No. S45~l Hole No.. 



Depth 



.ft (. 



.m) 



Figure 3. Size analysis and relative density of fine sand backfill material. 



11 




Figure 4. Test 2. Contours 
of deposited backfill 
material at end of test. 
Mine cavity submerged. 
Mine roof was lifted from 
the pillars by the pump 
pressure. (Preliminary 
test — distorted model 
scale) 



10 X 40 X 6 PILLAR BLOCKS 



Y//VA BACKFILL MATERIAL 

FILLED TO 900F LE-VEL 




Figure 5. Test 2. Pressure in the mine cavity resulting from slurry 
pumping caused the roof to raise above the pillars, allowing slurry to 
flow over the tops of pillars. Mine submerged and dipping 5 . 
(Distorted model scale) 



12 



-DOWNWARD 5° SLOPE 

U U U U L_ 




10' X 40' X 6' PILLAR BLOCKS 



Figure 6. Test 3. Contour map of backfill deposit at end of test. Mine cavity submerged. Mine roof was bolted tight to the pillars. 
(Distorted model scale) 



13 



i nfin 



r 

i 
L 







L 







Figure 7. Test 4. Contour 
map of backfill deposit at 
end of test. Mine cavity 
submerged. The pattern of 
deposit is typical for a 
small amount of backfill 
material. (Distorted model 
scale) 



10 x 40 X 6 PILLAR BLOCKS 




Figure 8. Test 4. The position of the pump intake was changed 
before test started, causing only a small deposit in the mine cavity 
during the test. Mine submerged and dipping 5 . (Distorted model 
scale) 



14 



n n n n! n 

f-DOWNWARD 5° SLOP 

J^U U WliU 




Figure 9. Test 5. 
Continuation of test 4. 
Contour map of deposit 
pattern at end of test. 
Mine cavity submerged. 
Compare deposit pattern 
with figure 7. (Distorted 
model scale) 



10 X 40 X 6 PILLAR BLOCKS 



Y////A BACKFILL MATERIAL 

FILLED TO ROOF LEVEL 




Figure 10. Test 5. This test was a continuation of test 4 in which 
the backfill deposit was allowed to continue. Mine submerged and 
dipping 5 . (Distorted model scale) 



15 



















































" 




















1 




nnnnnn 

-DOWNWARD 5° SLOPE 




i ii i r — 1 i i 

ST BREAKOUT CHANNEL 
EFORE TEST WAS 
STOPPED 



u u 



































r~ 




































_ 



p 



u u 



10 X 40 X 6 PILLAR BLOCKS 



V////A BACKFILL MATERIAL 

FILLED TO ROOF LEVEL 



Figure 11. Test 7. Contour map of backfill deposit at end of test. Mine submerged. Inside diameter of injection pipe was 
increased from 1/2 to 3/4 inch for this and later tests. (Model scale was undistorted for this and all later tests) 



16 





Figure 12. Test 7. Mine pillars and cavity were changed to give an ungistorted scale 
im : 48P f or tms test and all later tests. Mine submerged and dipping 5 . 



17 




10 X 40 X 6 PILLAR B 



LOCKS-^ / 



V////A BACKFILL MATERIAL 

FILLED TO ROOF LEVEL 



Figure 13. Test 8. Contour map at end of test. Mine submerged. Flow along the last breakout channel occurred at the end of the 
test. Full flow of the injection pipe down one channel caused equilibrium conditions for sediment transport in one channel to 
occur. Note that breakout channel flows upslope. 



18 





Figure 14. Test 8. Backfill material filled the mine cavity up to or near the roofline 
over a comparatively large area, see figure 13. The slurry concentration was 11.4 
percent by weight for this test. Mine submerged and dipping 5 . 





Figure 15. Test 9. First test in the series of three tests, 9—11. Blind entries were 
simulated by blocking corridors at various places. Mine submerged and dipping 5 . 



20 



L. 

r 1 



"in [ inn 

■ DOWNWARD 5° SLOPE 

n n n u lLj 

H 







"1 


















~"i n 
i 




L- 




! 
J 



















^J L_i 





L.J L 



J 

n 




u 

10' X 40' X 6' PILLAR BLOCKS- 



X^ 



I ! I 

FULL HEIGHT CORRIDOR BLOCKS 

I I I I I I I I 



Figure 16. Test 9. 
Location and pattern of 
blind entries simulated by 
solid blocks in the 
corridors. Contour map 
shows deposit pattern of 
backfill at the end of test 
9. 



| 2 BACKFILL MATERIAL 

FILLED TO ROOF LEVEL 




Figure 17. Test 11. Corridor 8 is pointed out in which the total 
injection flow was concentrated after backfill material filled to the 
roof around the injection hole. Mine submerged and dipping 5 . 
Last test in series of thre& tests, 9-1 1 . 



21 



nnn:' 

DOWNWARD 5° 



LAST BREAKOUT CHANNEL 
BEFORE TEST WAS STOPP 




10 X 40 X 6 PILLAR BLOCKS 



V////A BACKFILL MATERIAL 

FILLED TO ROOF LEVEL 



Figure 18. Test 11. Contours show deposit pattern of backfill material at the end of the test. The location and pattern of blind 
entries affect the general pattern of backfill ueposit. 



22 




10 X 40 X 6 PILLAR BLOCKS 



E-^S^Sl DEAD FLOOR AREA FORMED BY 
CORRIDOR BLOCKS. 

Y////A DEPOSITS SLICED AT 3 FOOT 
ELEVATION. 



Figure 19. Test 11. Deposits above the 3-foot contour elevation cover a large area around the injection hole. Overlapping 
crosshatching indicates entry of fill material into blind entries. 



23 




Pillars and simulated roof falls. 




Roof and openings for cavities. 

Figure 20. Test 17. Mine pillars and roof before tests simulating roof falls and 
cavities over the roof falls. 



24 




ROOF FALL CAVITY 
SIMULATED BY BOX 

MINE PILLAR 



-END OF PILLAR 



-RUBBLE PILE SIMULATED 
WITH WOODEN BLOCKS 
AND WEDGES 

SCHEMATIC SECTION A-A 



10 



u 



9 



n n 

SIMULATED 

I I I I 
ROCKFALLS 

u u 





3 






Ia 
1 


6 


A 


i 



PIEZOMETER TAPS 




10 X 40 X 6 PILLAR BLOCKS 

Figure 21 . Test 17. Cross section and layout for tests with roof falls and roof cavities over roof falls. 



25 





Figure 22. Test 17. Deposit pattern of backfill at the end of the test. Note backfill 
deposited in cavities over roof falls. 



26 



LEVEL MINE CAVITY 



nnnr 

*- ROOF FALL 

: u u u u 




10' X 40' X 6' PILLAR BLOCKS 



V////A BACKFILL MATERIAL 
FILLED TO TOP OF 
ROOF FALL CAVITY 



Figure 23. Test 17. Contours of deposit pattern at the end of the test. Last breakout channel flowed over a roof fall. 



27 




10 X 40 X 6 PILLAR BLOCKS 



V////A BACKFILL MATERIAL 
FILLED TO TOP OF 
ROOF FALL CAVITY 



Figure 24. Test 18. This test was a duplicate of test 17 except test 18 had a higher injection velocity and a higher slurry 
concentration. Note breakout channel in a corridor adjacent to a roc kf all. 



28 



n n n n n 




EPOSIT 



I 



u 



10 X 40 X 6 PILLAR BLOCKS 

Figure 25. Test 12. The first of a series of four tests, 12-15, made in a dry cavity with a level floor. Contour map shows backfill 
deposit pattern at the end of the test. 



29 





Figure 26. Test 12. The low height of deposit is a result of injection into a dry 
cavity with a level floor. 



30 




10 X 40 X 6 PILLAR BLOCKS 



Figure 27. Test 13. The second in the series of four tests with slurry injected into a dry cavity with a level floor. Contours show 
accumulated deposit pattern. 



31 





Figure 28. Test 13. Accumulated deposit after the second in the series of four tests 
shows shallow deposits and flat slopes on the surface of the deposited backfill 
material. 



32 




10' X 40' X 6' PILLAR BLOCKS 



Figure 29. Test 14. The backfill material continues to build up as the third in the series of four tests is completed in a dry cavity 
with a level floor. 



33 





Figure 30. Test 14. Deposit around the injection hole is uniform in a dry cavity 
with a level floor. 



34 




10' X 40' X 6' PILLAR BLOCKS 



Figure 31. Test 15. Contour map showing backfill deposit at the end of the series of four tests. Note the last breakout channel 
before the end of the test. 



35 





Figure 32. Test 15. Backfill material has built up near the roof close to the 
injection point at the end of the series of four tests in a dry mine with a level floor. 



36 



innnn 

EL MINE CA\ 

uuu 



LEVEL MINE CAVITY 




10' X 40' X 6' PILLAR BLOCKS 



Y////A BACKFILL MATERIAL 

FILLED TO ROOF LEVEL 



Figure 33. Test 16. Contour map of deposited fill material at end of test. Compare the deposit pattern on this level mine test 
with test 8 (fig. 13) in a sloping mine. 



37 





Figure 34. Test 16. the final breakout channel is pointed out on the photograph. 
Test was for a mine with a level floor and submerged. 



38 



7-1750(3-71) 

Bur«ou of Reclamation 

CONVERSION FACTORS-BRITISH TO METRIC UNITS OF MEASUREMENT 

The following conversion factors adopted by the Bureau of Reclamation are those published by the American 
Society for Testing and Materials (ASTM Metric Practice Guide, E 380-68) except that additional factors (*) 
commonly used in the Bureau have been added. Further discussion of definitions of quantities and units is given in 
the ASTM Metric Practice Guide. 

The metric units and conversion factors adopted by the ASTM are based on the "International System of Units" 
(designated SI for Systeme International d'Unites), fixed by the International Committee for Weights and 
Measures; this system is also known as the Giorgi or MKSA (meter-kilogram (mass)-second-ampere) system. This 
system has been adopted by the International Organization for Standardization in ISO Recommendation R-31. 

The metric technical unit of force is the kilogram-force; this is the force which, when applied to a body having a 
mass of 1 kg, gives it an acceleration of 9.80665 m/sec/sec, the standard acceleration of free fall toward the earth's 
center for sea level at 45 deg latitude. The metric unit of force in SI units is the newton (N), which is defined as 
that force which, when applied to a body having a mass of 1 kg, gives it an acceleration of 1 m/sec/sec. These units 
must be distinguished from the (inconstant) local weight of a body having a mass of 1 kg, that is, the weight of a 
body is that force with which a body is attracted to the earth and is equal to the mass of a body multiplied by the 
acceleration due to gravity. However, because it is general practice to use "pound" rather than the technically 
correct term "pound-force," the term "kilogram" (or derived mass unit) has been used in this guide instead of 
"kilogram-force" in expressing the conversion factors for forces. The newton unit of force will find increasing use, 
and is essential in SI units. 

Where approximate or nominal English units are used to express a value or range of values, the converted metric 
units in parentheses are also approximate or'nominal. Where precise English units are used, the converted metric 
units are expressed as equally significant values. 

Table I 

QUANTITIES AND UNITS OF SPACE 



Multiply 



By 



To obtain 



LENGTH 



Mil 

Inches .... 
Inches .... 

Feet 

Feet 

Feet 

Yards .... 
Miles (statute) 
Miles 

Square inches 
Square feet . 
Square feet . 
Square yards 

Acres 

Acres 

Acres 

Square miles 

Cubic inches 
Cubic feet . . 
Cubic yards . 



Fluid ounces (U.S.) 
Fluid ounces (U.S.) 
Liquid pints (U.S.) 
Liquid pints (U.S.) 
Quarts (U.S.) . . . 
Quarts (U.S.) . . . 
Gallons (U.S.) . . . 
Gallons (U.S.) . . . 
Gallons (U.S.) . . . 
Gallons (U.S.) . . . 
Gallons (U.K.) . . 
Gallon? (U.K.) . . 

Cubic feet 

Cubic yards .... 

Acre-feet 

Acre-feet 



25.4 (exactly) Micron 

25.4 (exactly) Millimeters 

2.54 (exactly)* Centimeters 

30.48 (exactly) . Centimeters 

0.3048 (exactly)* Meters 

0.0003048 (exactly)* Kilometers 

0.9144 (exactly) Meters 

1,609.344 (exactly)* Meters 

1.609344 (exactly) Kilometers 

AREA 

6.4516 (exactly) Square centimeters 

*929.03 Square centimeters 

0.092903 Square meters 

0.836127 Square meters 

*0.40469 Hectares 

*4,046.9 Square meters 

"0.0040469 Square kilometers 

2.58999 Square kilometers 

VOLUME 

16.3871 Cubic centimeters 

0.0283168 Cubic meters 

0.764555 Cubic meters 

CAPACITY 

29.5737 Cubic centimeters 

29.5729 Milliliters 

0.473179 Cubic decimeters 

0.473166 Liters 

*946.358 . . . . , Cubic centimeters 

'0.946331 Liters 

*3,785.43 Cubic centimeters 

3.78543 Cubic decimeters 

3.78533 Liters 

*0.00378543 Cubic*meters 

4.54609 Cubic decimeters 

4.54596 Liters 

28.3160 Liters 

*764.55 Liters 

*1,233.5 Cubic meters 

'1,233,500 Liters 



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